TESTING  FOR 
METALLURGICAL 


MINING    D£] 


D.  B.  Huntley 


Testing  for 
Metallurgical  Processes 


BY 

JAMES  A.  BARR,  B.  s.,  M.  E. 

Engineer  of  Construction  and  Maintenance  for  the 

Charleston,  S.  C.,  M.  &  M.  Company.     Formerly 

Instructor,  Michigan  College  of  Mines. 


FIRST    RDTION 


1910 

Published  by  the 
Mining  and  Scientific  Press,  San  Francisco, 

and 
The  Mining  Magazine,  London. 


COPYRIGHT,  1910 

BY 
DEWEY  PUBLISHING  COMPANY 


PREFACE 

This  book  is  based  upon  original  notes  and  ex- 
periments made  in  the  laboratories  of  the  Michigan 
College  of  Mines,  and  the  nucleus  of  the  material 
was  published  in  1909  in  pamphlet  form  under  the 
titles  'Metallurgical  Laboratory  Experiments'  and 
'Ore  Dressing  Laboratory  Experiments.'  These 
booklets  have  been  used  as  text-books  in  the  College. 
In  preparing  the  present  edition  the  text  has  been 
extensively  revised  and  enlarged.  Free  use  has  been 
made  of  standard  reference  books,  the  more  impor- 
tant of  which  are  listed  below.  I  am  under  especial 
obligations  to  S.  S.  Bruce,  professor  of  metallurgy 
and  ore  dressing  at  the  Michigan  College  of  Mines, 
for  advice  and  assistance  in  preparing  the  new  text, 
particularly  that  part  relating  to  slags  and  slag  cal- 
culations. 

JAMES  A.  BAKU. 

Mt.  Pleasant,  Tennessee, 
August  26,  1910. 


438547 


TABLE  OF  CONTENTS 


Page. 

Amalgamation 9 

Chlorination   14 

General    14 

The  Oxidizing  Roast 16 

Solution  of  the  Gold 18 

Precipitation    19 

Procedure  for  the  Test 19 

Cyanidation    24 

Preliminary  Investigations 24 

Chemical  Tests 26 

1.  Consumption   of   Cyanide    26 

2.  Total  Acidity    27 

3.  Best  Strength  of  Solution 28 

4.  Fineness  of  Grinding 32 

5.  Time  of  Treatment 33 

6.  Rate  of  Percolation 35 

7.  Sizing  Test 40 

8.  Sliming  Test  . 42 

9.  Decantation  Tests   44 

10.  Determination  of  the  Causes  of  Consumption . .  45 

General  Remarks  50 

Causes  for  the  Non-Extraction 50 

Cyanide  Poisoning  51 

Reactions  in  the  Cyanide  Process 51 

In  the  Preliminary  Treatment 51 

In  the  Solution  Tanks 51 

In  the  Air 52 

Substances  in  Solution 52 

In  Precipitating  Tanks  and  Zinc  Boxes 52 

Cyanide  Tests  on  Lots  of  30  Pounds  and  More  of  Ore.  53 

Double  Treatment 55 

Precipitation  by  Means  of  Zinc '. 56 

Treatment  of  Concentrate 61 

Sliming  Tests  for  5  to  20  pounds  of  Ore. .                 .  62 


6  CONTENTS 

Page. 

Cyanidation  of  Silver  Ores    64 

Chloridizing  Roast  of  Silver  Ores 66 

Pan  Amalgamation  for  Silver  Ores 71 

Pot  Roasting 76 

Parkes  Process 80 

Ziervogel  Process  or  the  Sulphatizing  Roasting  of  Copper 

Matte 86 

The  General  Process  86 

Methods  of  Analysis 90 

Re-sulphatizing  the  Metallic  Silver 91 

Lixiviation  of  Copper  Ores 94 

Electrolytic  Processes   97 

General  Statement  97 

Electrolytic  Refining  of  Copper 99 

Electrolytic  Refining  of  Lead— Betts  Process 104 

Zinc  Smelting 109 

Furnace  Test  113 

Methods  of  Calculation  for  Heat-Balance  Sheet 118 

Approximate  Determination  of  the  Calorific  Power  of  a 

Fuel 125 

Proximate  Analysis 125 

Malher   Formula    127 

Pyrometry     128 

Retorting 137 

Refining  of  Bullion 140 

Concentration   Tests    143 

Preliminary  Observations    143 

Screen  Sizing  Tests  145 

Wet  Concentration    146 

Magnetic  Separation    150 

Electrostatic    Separation    151 

Oil   Flotation    153 

Determination  of  the  Suitability  of  an  Ore  for  Smelting.  .155 
Schedule  for  Dry  Ores,  Concentrate,  and  Tailing.  .156 

Schedule  for  Ores  and  Concentrate 159 

Physical  Properties  of  Slags 160 

Smelting  Properties  of  Copper,  Lead,  Gold,  and  Silver 


CONTENTS  7 

Page. 

Ores 164 

Lead  Slags   165 

Slag  Types 166 

Remarks 167 

Slag    Constituents    167 

General  Data   167 

Effect  of  Slag  Constituents 168 

General  Remarks 169 

Copper  Slags    173 

Constituent  Limits  for  Copper  Slags  and  Furnace 

Data    176 

General  Remarks 178 

Hoffman's  Slag  Tables   184 

Calculation  of  a  Silver-Lead  Blast-Purnace  Charge.  .186 

Calculations  of  a  Copper  Blast-Furnace  Charge 192 

Application  of  Slag  Data 195 

Cost  Data 198 

General 198 

Ore  Treatment  198 

Amalgamation  and  Concentration    198 

Oil  Flotation    200 

Chlorination   200 

General    200 

Cyanidation    201 

Cost  of  Hyposulphite  Lixiviation 203 

Cost  of  Huntington-Heberlein  Pot-Roasting 203 

Cost  of  Lead  Smelting 204 

Refining  Lead  Bullion    204 

Cost  of  Copper  Smelting 205 

Cost  of  Copper  Smelting  Works 206 

Cost  of  Copper  Leaching 206 

Cost  of  Electrolytic  Copper  Refining 206 

Betts  Electrolytic  Process  207 

Cost  of  Zinc  Smelting 207 

Cost  of  Zinc  Smelters. .  .  .208 


AMALGAMATION 

Amalgamation  is  based  on  the  affinity  of  mercury 
for  gold  in  the  metallic  state  and  consists,  briefly, 
in  crushing  the  ore  to  the  necessary  degree  of  fine- 
ness to  release  the  metallic  particles  and  bringing 
them  in  contact  with  mercury  with  which  they  be- 
come alloyed,  or  amalgamated..  For  gold  ores  in 
which  the  precious  metal  is  in  a  condition  to  amal- 
gamate with  mercury,  that  is,  '  free  milling '  ores,  the 
amalgamation  process  offers  advantages  over  all 
others  in  simplicity  and  economy. 

In  considering  the  treatment  of  gold-bearing  ore, 
it  is  usually  necessary  to  obtain  sufficient  data  to 
fully  answer  the  following  questions:  (1)  What  is 
the  value  of  the  ore?  (2)  Is  the  ore -free  milling?  If 
so,  what  percentage  of  the  gold  can  be  amalgamated? 
(3)  What  value  has  the  tailing  after  amalgamation? 
Can  it  be  further  treated  at  a  profit?  (4)  What  per- 
centage of  concentrate  does  the  original  ore  carry? 
Can  the  ore  be  profitably  concentrated  or  could  the 
tailing  from  the  amalgamating  plates  be  better 
treated  by  other  methods? 

These  questions  can  be  best  answered  by  making 
a  run  of  10  to  20  tons  of  the  ore  in  a  regular  mill 
fully  equipped  for  amalgamation  and  concentration. 
Before  making  a  trial  on  such  a  large  and  necessarily 
expensive  scale,  the  following  preliminary  tests 
should  be  made. 

Test. — Take  a  500-gram  sample  of  the  ore  and 
crush  gradually  so  as  to  pass  a  40-mesh  screen.  Es- 


10  TESTING  FOR  PROCESSES 

pecial  care  should  be  observed  in  regard  to  any  resi- 
due or  metallic  particles  left  on  the  sieve.  If  any 
such  particles  are  found  they  should  be  carefully 
saved,  separately  assayed,  and  the  assay  value  of  the 
final  sample  corrected  accordingly.  Carefully  mix 
the  40-mesh  ore,  take  a  50-gram  sample  for  assaying, 
crush  through  120  mesh  and  determine  the  gold,  sil- 
ver, base-metal  percentages,  and  amount  insoluble. 

Free  Milling  Test.— Take  a  5  A.  T.  sample  of  the 
40-mesh  ore,  introduce  carefully  into  a  wide-mouthed 
bottle,  add  clean  mercury  equivalent  to  10%  of  the 
weight  of  the  ore  taken,  with  sufficient  water  to 
make  the  whole  into  a  thick  pulp.  Cork  up  the 
bottle  securely  and  rotate  from  2  to  3  hours.  Sepa- 
rate the  mercury  from  the  tailing  by  a  hydraulic 
classifier,  Fig.  1,  or  by  panning  with  an  ordinary 
gold  pan.  Save  all  the  water,  concentrate,  tailing, 
and  slime,  allow  to  stand  until  the  supernatant  water 
becomes  clear,  then  decant  the  water  using  care  not 
to  lose  any  of  the  slime.  Dry  the  remaining  tailing, 
mix  well,  sample  and  assay  for  gold.  The  differ- 
ence in  the  assay  of  the  original  ore  and  the  amount 
found  above  will  equal  the  gold  amalgamated.  Cal- 
culate the  percentage  of  gold  recovered.  The  gold 
may  be  recovered  from  the  mercury  as  given  under 
i  Retorting. ' 

If  any  minerals  suitable  for  saving  by  concentra- 
tion are  present  in  the  ore,  the  following  tests  should 
be  made.  Take  a  100-gram  sample  of  the  tailing 
from  the  amalgamation  test  and  carefully  pan  down 
so  as  to  save  all  the  heavy  minerals  such  as  pyrite. 
A  larger  sample  may  be  taken  for  this  test  if  the 


AMALGAMATION 


11 


result  of  the  concentration  does  not  give  a  sample 
large  enough  for  assaying  purposes.     For  this  con- 


n 


Fig.  1.     HYDRAULIC  CLASSIFIER 

centration,  the  ordinary  miner's  gold  pan  is  used. 
(Fig.   2.)      The   weighed  sample   of  ore   should  be 


12 


TESTING  FOR  PROCESSES 


placed  in  the  pan  and  thoroughly  softened  with 
water;  the  pan  is  then  shaken  sideways  and  in  a 
circular  manner  to  give  heavy  particles  an  oppor- 


Fig.  2.     MINER'S  GOLD  PAN 

tunity  to  settle;  water  is  flowed  across  the  top,  re- 
moving the  overlying  layer  of  waste.  The  shaking 
and  flowing  are  repeated  until  the  concentrate  is 
sufficiently  cleaned  from  gangue.  Do  not  try  to  pan 
off  the  last  traces  of  the  gangue,  as  some  of  the  con- 
centrate will  pass  over  into  the  tailing.  Pan  the 
tailing  again  to  save  any  concentrate  that  may  have 
been  lost  in  the  first  operation.  Dry  and  weigh  both 
the  concentrate  and  the  tailing.  Assay  the  tailing 
for  gold  and  silver.  If  the  value  of  the  gold  and 
silver  in  the  tailing  is  sufficient,  their  further  treat- 
ment by  cyanidation  should  be  considered.  Assay 
the  concentrate  for  gold,  silver,  base  metals,  and  in- 
soluble. Calculate  the  percentage  of  concentrate  in 


AMALGAMATION  13 

the  ore ;  also  how  many  tons  of  ore  will  be  required 
to  produce  one  ton  of  concentrate.* 

Make  out  the  following  report  which  will  serve 
as  a  check  on  the  work  : 

150  grams  of  ore  amalgamated;  gold  content  as  per  assay 
=  0.0250  grams. 

Grams.  Percent. 

Gold  recovered,  calculated  by  difference.        0.020  80.0 
Gold  recovered  from  150  grams  of  ore  by 

concentration    0.0040  16.0 

Gold  in  tailing 0.0005  2.0 

Gold  unaccounted  for  (by  difference) ...        0.0005  2.0 

Total    0.0250         100.0 

Suppose  the  charge  for  amalgamation  alone  is 
$0.50  per  ton,  for  concentration  and  amalgamation 
together  is  $1  per  ton,  and  for  concentration  is  $0.75 
per  ton.  Allow  $2.50  for  mining  and  $1  for  freight 
charges  to  the  smelter.  With  these  costs,  the  data 
previously  obtained,  and  a  smelter  schedule,  figure 
the  profit  and  loss  in  treating  the  ore : 

1.  By  shipping  direct  to  smelter. 

2.  By  amalgamation  and  wasting  tailing.f 

3.  By  amalgamation,  concentration,  and  shipping 
concentrate  to  smelter. 


wt.  concentrate 

•Percentage  concentrate  =   X  100.    If  an 

wt.  ore  taken 

ore  contains  4%  of  concentrate,  25  ton  of  ore  will  concen- 
trate into  1,  for  100:4: :  X  :1. 

flf  sufficient  gold  was  not  recovered  by  amalgamation  to 
pay  the  cost  of  amalgamating  after  crushing  ($0.20),  that 
process  would  probably  be  omitted  in  the  treatment  of 
the  ore. 


14  TESTING  FOR  PROCESSES 

Include  all  information  and  useful  data  obtained, 
in  a  concise  report  with  remarks  as  to  the  best  prob- 
able method  of  treating  the  ore  in  question. 


CHLOBINATION 

GENERAL 

The  chlorination  process,  which  is  based  upon  the 
solubility  of  gold  in  chlorine  water,  offers  an  avail- 
able means  for  extracting  the  gold  from  some  classes 
of  ores  that  are  not  amenable  to  copper-plate  or  pan 
amalgamation,  or  to  the  economical  treatment  by 
cyanidation,  lixiviation,  or  smelting.  With  some 
classes  of  ores,  the  chlorination  process  may  give 
better  results  than  treatment  with  cyanide.  As  a 
rule,  the  cyanidation  process  is  better  adapted  for 
the  treatment  of  low-grade  ores,  while  for  the  treat- 
ment of  many  higher-grade  refractory  ores,  chlorina- 
tion gives  better  results.  The  treatment  of  such 
material  by  chlorination  involves  the  roasting  of  the 
ore  which  is  usually  the  most  expensive  part  of  the 
operation. 

Ores,  in  which  the  gangue  is  hydrated  oxide  of 
iron,  are  extremely  difficult  to  amalgamate,  either 
by  plate  or  pan  amalgamation,  as  the  gold  is  not 
only  very  finely  divided,  but  the  pulp  is  usually 
excessively  slimy,  enveloping  the  particles  of  gold 
and  forming  an  almost  impervious  coating  on  the 
amalgamation  plates.  For  this  class  of  ores,  barrel 
chlorination  gives  very  satisfactory  results.  For 
extracting  gold  from  some  classes  of  concentrate, 
especially  sulphides,  and  also  for  the  treatment  of 
arsenical  pyrites,  the  chlorination  process  has  been 
successfully  applied.  Some  classes  of  pyritic  ores 


16  TESTING  FOR  PROCESSES 

contain  an  amount  of  acid  making  the  economical 
treatment  by  the  cyanide  process  prohibitive,  but 
the  same  ores  yield  excellent  results  with  chlorina- 
tion.  The  chlorination  process  can  be  advantage- 
ously utilized  for  treating  ores  containing  particles 
of  so-called  'rusty  gold'  or  particles  coated  with 
other  substances  that  make  them  refractory  to  amal- 
gamation, and  which  at  the  same  time  are  too  coarse 
for  cyanidation.  Ores  containing  much  lime  or  mag- 
nesia can  not  be  successfully  treated  by  the  chlorina- 
tion process  as  there  would  be  an  excessive  con- 
sumption of  chlorine  by  these  constituents. 

THE  OXIDIZING  ROAST 

Since  roasting  is  the  primary  step  in  the  process, 
it  will  be  described  in  detail  first.  The  object  of 
roasting  is  to  burn  the  sulphur  of  the  ore  to  sulphur 
dioxide,  SO2,  which  passes  away  as  a  gas,  and  to 
change  into  an  oxide  the  metal  with  which  this  sul- 
phur was  combined.  It  is  not  possible  in  practical 
work  to  drive  off  all  the  sulphur  or  to  change  all 
the  metal  into  oxide.  Moreover  it  takes  a  dispro- 
portionate amount  of  time  and  heat  to  remove  the 
last  per  cent  of  sulphur.  In  the  case  of  roasting  an 
ore  preparatory  to  the  smelting  process,  it  is  not 
desirable  to  oxidize  all  the  sulphur  present  in  the 
ore,  but  in  roasting  an  ore  preparatory  to  chlorina- 
tion or  cyanidation,  it  is  sought  to  remove  most  or 
all. of  the  sulphur,  to  have  the  least  amount  of  sul- 
phates, and  as  much  of  the  metal  in  the  oxide  form 
as  possible. 

The  requirements  for  good  roasting  are: 


18  TESTING  FOR  PROCESSES 

1.  The  ore  should  be  in  a  finely  divided  form  so 
that  the  oxygen  of  the  air  can  come  into  contact 
with    each    particle    of   sulphide    which    should    be 
stirred  frequently. 

2.  The  bed  of  ore  should  be  thin. 

3.  A  current  of  air  should  constantly  sweep  away 
the  gases  resulting  from  the  roasting  reactions,  and 
bring  a  fresh  supply  of  oxygen. 

4.  The  ore  should  be  sufficiently  heated  to  bring 
about  the  reactions  desired,  and  yet  not  enough  to 
melt  the  easily  fusible  sulphides.     A  dull  red  heat 
answers  the  above  requirements. 

5.  The  gases  used  for  heating  the  ore  must  be 
highly  oxidizing.     Since  the  surface  of  the  ore  is 
experiencing  rapid  oxidation,  it  is  plain,  that  if  the 
expanse  of  surface  can  be  increased  without  enlarg- 
ing the  hearth  area,  a  larger  quantity  of  ore  can  be 
roasted  in  a  furnace  of  the  same  size.    The  increase 
may  be  effected  by  furrowing  the  ore  with  the  aid 
of  rakes  or  small  ploughs,  so  that,  instead  of  lying 
in  a  smooth  horizontal  bed,  the  ore  shall  be  in  a 
series  of  ridges  and  furrows  like  a  plowed  field. 

SOLUTION  OF  THE  GOLD 

The  solution  of  the  gold  is  effected  either  in  tanks 
or  in  revolving  barrels. 

In  the  former  method,  which  is  known  as  the 
'Plattner  process',  the  roasted  ore,  slightly  moist- 
ened with  water,  is  charged  into  acid-proof  tanks 
supplied  with  tightly  fitting  covers  and  a  false  bot- 
tom or  filter-bed.  Upon  receiving  the  charge  of  ore, 
the  covers  are  replaced  and  hermetically  sealed,  after 


CHLORINATION  19 

"which  chlorine  gas  is  forced  into  the  moist  ore,  and 
converts  the  gold  into  the  soluble  tri-chloride  accord- 
ing to  the  reaction : 

Au  +  3C1  =  AuCl3. 

The  gold  chloride  formed  is  leached  out  with 
water.  The  gold-bearing  solution  is  run  into  settling 
tanks  where  the  gold  is  precipitated  and  collected. 

In  the  second  method,  known  as  the  *  barrel  pro- 
cess', the  moist  ore  is  charged  into  lead-lined  barrels 
fitted  with  a  filter-bottom.  The  chlorine  is  generated 
directly  in  the  barrel  by  sulphuric  acid  and  bleaching 
powder  according  to  the  reaction: 

H2S04  +  CaOCl2  =  CaS04  +  H20  +  2C1. 

During  the  solution  process,  the  barrel  is  contin- 
ually rotated.    The  auric  chloride  is  leached  out  with 
water  as  in  the  previously  described  process. 
PRECIPITATION 

The  precipitation  may  be  effected  either  by  the 
addition  of  ferrous  sulphate  or  hydrogen  sulphide. 
The  latter  method  is  more  generally  used.  The  ex- 
cess of  chlorine  gas  is  first  eliminated  by  sulphur 
dioxide : 

2C1  +  S02  +  2H20  =  H2S04  +  2HC1. 

The  gold  is  precipitated  by  hydrogen  sulphide. 
2AuCl3  +  3H2S  =  Au2S3  +  6HCL 

The  resulting  precipitate  is  collected  in  a  filter- 
press,  dried  and  refined. 

PROCEDURE  FOR  THE  TEST 
Take  a  500-gram  sample  of  ore  and  assay  for  gold, 
silver,  base  metals,  and  sulphur. 


20  TESTING  FOR  PROCESSES 

The  ore  should  then  be  crushed  so  as  to  pass 
through  a  40-mesh  screen,  carefully  weighed  and 
placed  in  an  iron  roasting  dish  in  a  thin,  even  layer. 
The  success  of  this  test  depends  upon  the  roast  being 
'dead';  that  is,  the  ore  should  be  roasted  until  prac- 
tically all  the  sulphur  is  driven  off. 

The  roasting  dish,  containing  the  ore,  should  then 
be  placed  in  the  oven  or  reverberatory  furnace.  The 
temperature  of  the  furnace  should  be  gradually  in- 
creased until  the  sulphur  is  seen  to  take  fire  and 
burn  with  a  blue  flickering  flame,  at  which  point  the 
temperature  is  held  stationary  until  the  sulphur 
ceases  to  burn  of  its  own  accord.  Rabble  the  charge 
frequently  so  as  to  prevent  caking.  Carry  on  the 
roast  with  continually  increasing  temperature  up  to 
the  maximum  limit  allowed  by  the  nature  of  the  ore, 
until  the  sulphur  content  does  not  exceed  0.2  or  0.3% 
(to  be  determined  by  analysis) .  This  roast  will  take 
from  4  to  5  hours.  During  the  roast,  endeavor  to 
keep  the  mechanical  losses  of  the  ore  as  low  as  pos- 
sible. 

The  decomposition  of  copper  and  iron  sulphates 
is  complete  at  850°  C.,  but  this  is  by  no  means  an 
instantaneous  action.  Any  ferrous  sulphate  left  in 
the  roast  would  be  oxidized  to  the  higher  form,  which 
action  is  not  harmful  except  to  consume  chlorine. 
The  reaction  is, 

6FeS04  +  3C12  =  2Fe2(SOJ3  +  Fe2Cl6. 

Undecomposed  sulphides  and  charcoal  are  detri- 
mental, as  they  precipitate  the  gold  prematurely  ac- 
cording to  the  following  reactions: 


CHLORINATION  21 

2AuCl3  +  3CuS  =  Au2S3  +  3CuCl2,  or 
8AuCls  +  12H2O  =  Au  +  24HC1  +  3CuS04. 
4AuCl3  +  3C  +  6H20  =  4Au  +  12HC1  +  3C02. 

After  the  roast  has  been  completed  and  the  ore 
cooled,  weigh  carefully  and  determine  the  loss  in 
weight  due  to  the  conversion  of  the  sulphides  into 
oxides,  in  addition  to  the  mechanical  losses.  The 
latter  will  be  small  enough  to  neglect  if  the  experi- 
ment has  been  carefully  performed.  Break  up  any 
lumps  in  the  roast  and  pass  through  a  30-mesh 
screen.  Sample  and  assay  for  gold  and  sulphur. 
Calculate  the  total  amount  of  gold  and  sulphur  in 
the  roasted  sample.  Having  already  determined  the 
total  amounts  in  the  original  sample  taken,  the  losses 
of  each  are  found  by  difference,  from  which  the  per- 
centages of  loss  are  determined.* 

In  making  the  chlorination  test  introduce  3  A.  T. 
of  the  roast  ore  into  a  strong  bottle  together  with 
the  following  charge:  bleaching  power  (35%  Cl), 
6%f ;  sulphuric  acid,  commercial,  12%;  water  as  re- 
quired. 

The  quantity  of  bleach  required  varies  with  the 
nature  of  the  ore  and  the  economical  amount  can 
only  be  found  by  trial.  The  quantity  of  acid  re- 
quired for  a  given  amount  of  bleach  is  variable. 
Theoretically,  7  parts  of  bleaching  powder  require 
6  parts  of  acid,  but  in  practice  this  proportion  is 
seldom  used,  the  proper  amount  being  determined 


'Loss  S  in  roast 

X  100  =  per  cent  loss  of  sulphur. 


Wt.  in  original 
fSix  per  cent  of  the  weight  of  ore  taken. 


22  TESTING  FOR  PROCESSES 

by  the  character  of  the  ore.  An  amount  of  acid  in 
excess  of  the  theoretical  requirement  is  recommended, 
otherwise  sulphate  of  lime  is  liable  to  be  precipitated 
with  the  gold  during  subsequent  operations. 

A  method  of  charging  barrels  in  practice  is  to 
add  first  the  requisite  amount  of  water,  then  the 
sulphuric  acid,  after  which  the  barrel  is  charged  with 
the  roasted  and  pulverized  ore,  and  lastly  the  bleach- 
ing powder.  By  this  method  the  chlorine  is  not 
generated  until  the  barrel  is  sealed  and  rotated.  In 
the  laboratory,  to  the  roasted  ore  add  enough  water 
to  make  the  whole  charge  into  a  thick  pulp.  The 
bleach,  wrapped  in  a  piece  of  filter  paper,  is  added, 
the  cork  put  tightly  in  place  and  securely  tied  down. 
Wrap  the  bottle  in  some  burlap  to  prevent  trouble 
in  case  the  bottle  bursts.  Rotate  the  bottle  contain- 
ing the  charge  from  4  to  6  hours,  at  the  end  of  which 
time  it  may  be  again  opened,  when  the  odor  of 
chlorine  should  be  very  noticeable  if  sufficient  chem- 
icals have  been  added  to  the  charge. 

Leach  the  charge  by  transferring  to  a  large  filter 
paper  held  in  a  ribbed  funnel  and  adding  wash- 
water  until  all  the  gold  chloride  has  been  washed 
out.  The  gold  may  be  recovered  from  the  solution 
by  precipitating  with  hydrogen  sulphide,  filtering 
off  the  resulting  precipitate,  scorifying  with  test  lead 
and  cupelling  for  gold.  Dry  the  leached  ore,  sample 
and  assay  for  gold.  The  amount  of  gold  extracted 
is  calculated  from  the  difference  in  the  assays  of  the 
roast  ore  and  the  leached  product. 


CHLORINATION  23 

Make  out  a  report  in  the  following  manner: 


Weight 

Silver 

Gold 

Sul- 
phur 

Re- 
marks 

Original    ore  

Roasted  ore.  .  .  . 

Losses  in  roast.  . 

Leached    ore  

Extraction 

Actual    recovery.  . 

Prepare  an  itemized  statement  of  costs,  including 
profit  and  loss.  Show  profit  by  shipping  ore  or  con- 
centrate direct  to  smelter. 


CYANIDATION 

In  making  tests  on  gold  and  silver  ores  to  deter- 
mine if  they  are  amenable  to  treatment  by  the  cya- 
nide process,  the  greatest  care,  accuracy  of  manipula- 
tion, and  good  judgment  are  necessary.  A  correct 
and  representative  sample  is  the  first  and  most  im- 
portant consideration.  Among  other  points  to  be 
carefully  investigated  are  the  physical  and  chemical 
states  in  which  the  gold  and  silver  exist ;  the  proper 
methods  of  grinding  the  ore;  the  fineness  of  grind- 
ing ;  the  presence  of  minerals  suitable  for  concentra- 
tion ;  the  solubility  of  gold  and  silver  in  the  cyanide 
solution;  the  consumption  of  chemicals;  the  precip- 
itation of  the  dissolved  metals.  Throughout  all  the 
investigations  the  question  of  profit  and  loss  must 
be  kept  constantly  in  view.  As  a  rule  the  experi- 
mental work  on  a  small  scale,  such  as  in  bottle  tests, 
should  not  be  taken  as  conclusive  nor  the  data  used 
to  construct  a  plant  for  the  commercial  treatment 
of  the  ores.  They  should  rather  serve  as  an  aid  to 
larger  tests  on  the  ore  in  lots  of  5  tons  or  more. 
Generally  speaking,  the  percentage  of  extraction  and 
the  consumption  of  cyanide  will  be  larger  in  the 
laboratory  tests  than  in  a  regular  mill-run. 

PRELIMINARY  INVESTIGATIONS 
Take  a  sample  of  30  to  40  pounds  of  the  ore,  from 
which  take  a  smaller  sample  of  2  pounds  for  the  fol- 
lowing tests. 

1.     Make  check  assays  for  gold  and  silver  and  thus 


CYANIDATION  25 

establish  the  value  of  the  ore.  If  below  $5  per  ton 
gold  the  conditions  must  be  very  favorable  to  permit 
cyaniding  the  ore. 

2.  Examine  the  ore  under  a  low-power  microscope 
and  determine  as  far  as  practicable  the  physical 
state  of  the  gold  and  silver  and  the  associated  min- 
erals. If  the  gold  appears  to  have  a  smooth  even 
surface,  or  appears  in  well-defined  crystals,  it  may 
be  taken  that  the  cyanide  will,  in  all  probability, 
dissolve  only  a  small  portion,  and  amalgamation  will 
be  necessary  if  these  particles  are  to  be  extracted. 
If,  on  the  other  hand,  the  gold  is  spongy  and  honey- 
combed in  appearance,  finely  divided,  or  invisible,  it 
is  possible  that  it  may  be  treated  direct  by  cyanide. 
Then,  again,  the  matrix  may  be  of  such  a  nature 
that  crushing  produces  an  excessive  quantity  of 
slime,  as,  for  instance,  when  an  ore  is  clayey  or 
consists  of  certain  shistose  rocks,  in  which  case  wet 
crushing  might  produce,  for  the  most  part,  a  product 
impervious  to  the  solution,  while  with  dry  crushing 
leaching  might  become  practicable  with  a  larger  per- 
centage of  ore.  The  ore  might  be  mainly  quartz 
and  then  either  wet  or  dry  crushing  could  be  adopted. 
The  decision  would  be  influenced  by  the  local  con- 
ditions and  the  physical  state  in  which  the  gold 
exists.  When  silver  is  present  in  chemical  combina- 
tion the  microscope  will  often  show  which  compound 
is  acted  on  by  the  solution  and  which  is  not.  This 
information  may  direct  in  an  important  manner  the 
treatment  method. 

The  presence  of  metallic  mineral  matter,  such  as 
sulphides  of  iron,  copper,  antimony,  or  zinc,  arsenides 


26  TESTING  FOR  PROCESSES 

and  tellurides,  should  always  be  determined  by 
microscope.  Their  systems  of  crystallization  and  the 
particular  minerals  present  should  be  identified,  as 
far  as  practicable,  in  every  unknown  ore.  This  ex- 
amination may  lead  to  important  results  bearing  on 
the  method  of  treatment  and  save  much  time  in 
experimenting.  In  cases  of  failure  such  an  examina- 
tion may  indicate  the  cause  and  methods  of  over- 
coming the  difficulty.  The  microscope  will  assist  in 
determining  whether  the  coarse  particles  of  gold  are 
amalgamable  before  or  after  cyanidation. 

CHEMICAL  TESTS 

1.  CONSUMPTION  OF  CYANIDE 
Introduce  20  grams  of  ore,  treated  with  a  sufficient 
quantity  of  alkali,  if  necessary,  as  found  in  (1),  into 
a  glass-stoppered  bottle ;  add  40  c.c.  of  KCN  solution 
of  known  strength,  agitate  for  20  minutes;  filter; 
measure  off  20  c.c.  of  the  filtrate  and  titrate  for  un- 
decomposed  KCN.  The  difference  in  the  amount  of 
KCN  in  20  c.c.  of  the  stock  solution  and  the  amount 
found  above,  gives  the  amount  of  KCN  consumed  by 
10  grams  of  ore.  Compute  in  terms  of  pounds  of 
KCN  per  ton  of  ore.*  If  not  over  4  pounds  of  KCN 
per  ton,  for  an  ordinary  grade  of  ore,  the  succeeding 
tests  may  be  tried. 

Tests  (1)  and  (2)  may  often  be  combined  to  ad- 
vantage as  follows:  Weigh  out  4  separate  20-gram 
samples  of  pulp,  place  in  glass-stoppered  bottles ;  add 


*Method  of  calculation. — Suppose  10  grams  of  ore  con- 
sumes 0.01  grams  of  KCN;  which  is  0.1%  of  the  weight  of 
ore,  0.1%  X  2000  =  2  Ib.  KCN  per  ton  of  ore. 


CYANIDATION  27 

fresh-slaked,  pure  lime,  at  the  rate  of  5,  10,  15,  and 
20  pounds  to  the  ton  of  ore,  and  then  40  c.c.  of  KCN 
solution  of  known  strength,  to  each  bottle ;  place  on 
the  agitator  for  20  minutes,  filter,  and  determine  the 
cyanide  consumption  as  given  above.  If  a  high  con- 
sumption of  cyanide  is  shown  even  with  15  to  20 
pounds  of  lime  per  ton,  see  if  the  soluble  salts  can 
be  removed  by  a  preliminary  water  wash,  using  3 
washes,  each  double  the  volume  of  the  ore. 

2.   .TOTAL  ACIDITY 

Gold  and  silver  ores  usually  contain  acid  products 
as  a  result  of  oxidation  of  sulphides  and  arsenides, 
that  are  soluble  in  water  and  have  the  power  of  neu- 
tralizing free  alkali.  In  some  cases  it  may  be  profita- 
ble to  give  the  ore  a  preliminary  wash  with  water 
to  get  rid  of  most  of  the  acid.  Ordinarily  the  acid 
is  neutralized  by  adding  an  alkali. 

Test. — Agitate  20  grams  of  the  pulp  for  20  minutes 
in  a  well-stoppered  bottle,  with  40  c.c.  of  tenth  nor- 
mal sodium  hydroxide  solution,  which  should  be  an 
excess  of  alkali;  filter  off  one-half  of  the  solution 
added,  and  titrate  the  free  alkali  with  tenth  normal 
acid.  Report  the  amount  of  lime  necessary  to  neu- 
tralize a  ton  of  ore.*  The  consumption  found  by 
the  above  titration  will  be  for  10  grams  of  ore. 

*Method  of  calculation. — 10  c.c.  of  acid  required  for  titra- 
tion; 20  c.c.  NaOH  contains  (a)  grams  NaOH;  10  c.c.  10/N 
acid  =  10  c.c.  10/N  alkali;  10  c.c.  10/N  NaOH  contains  (b) 
grams  NaOH;  a — b  =  grams  NaOH  consumed  by  10  grams 
of  ore. 

Then  10  (a — b)  =NaOH  consumed  by  100  gram  of  ore  or 
10  (a — b)%  X  2000  =  consumption  NaOH  per  ton  of  ore, 
which  is  converted  into  terms  of  CaO. 


28  TESTING  FOR  PROCESSES 

3.     BEST  STRENGTH  OF  SOLUTION 

Take  3  four-ounce,  wide-mouthed  glass  bottles, 
place  1  A.  T.  of  pulp  in  each  ;  then  add  60  c.c.  of  KCN 
solution  of  the  following  strength,  respectively  : 

To  No.  1  add  0.05%  KCN 
«     «     2     "     0.25"      " 


Before  adding  the  KCN  solution,  thoroughly  mix 
the  proper  amount  of  neutralizer  as  found  in  (1). 
Allow  to  stand  for  48  hours  with  occasional  shakings. 
Filter  off  equal  portions  of  the  solution  and  titrate 
for  undecomposed  KCN,  from  which  compute  the 
consumption  of  KCN  per  ton  of  ore.  Thoroughly 
wash,  dry,  and  assay  the  tailing  for  gold  and  silver. 
From  the  assay  of  the  original  pulp  and  the  assays 
of  the  tailing,  compute  the  percentages  of  extraction 
in  each  case. 

Physical  Effect  of  Strong  and  Weak  Solutions.  —  In 
practical  work,  the  strength  of  solution  was  found, 
in  some  cases,  to  have  a  marked  physical  effect  on 
the  dissolving  of  gold  and  silver.  This  is  most  no- 
ticeable in  ores  containing  a  preponderance  of  silver 
over  gold  in  the  presence  of  sulphides.  Thus  it  was 
found  when  employing  a  weak  solution  of  0.15% 
strength  KCN  a  maximum  extraction  of  66.6%  silver 
was  reached,  and  even  when  the  solution  was  after- 
ward strengthened  to  0.50%  the  additional  extrac- 
tion was  very  small.  When,  however,  the  first  solu- 
tion applied  was  0.50%  strength  and  afterward 
weaker  solutions  were  used,  the  extraction  was  over 
90%.  It  is  surmised  that  the  weak  solution  produced 


CYANIDATION  29 

a  hard,  insoluble  film  of  sulphide  over  the  surface 
of  a  portion  of  the  metallic  particles  which  prevented 
even  the  strong  solution  from  attacking  the  silver, 
whereas,  when  the  strong  solution  was  used  at  first, 
the  sulphide  adhered  as  a  loose  slimy  deposit,  which 
did  not  prevent  the  weaker  solution  from  afterward 
attacking  the  metallic  silver.  A  hot  solution  pro- 
duces the  same  slimy  deposit  even  when  weak. 

Comparative  Solubility  of  Metals  and  Minerals  in  Cya- 
nide Solutions. — It  has  been  found  by  experiment  that 
the  rate  of  dissolution  of  gold  and  silver  varies  with 
the  strength  of  solution,  being  small  for  strong  solu- 
tions and  increasing  as  the  solution  becomes  weaker 
until  a  maximum  at  0.25%  strength  of  KCN  is 
reached,  and  then  again  diminishing.  These  experi- 
ments were  performed  by  Maclaurin  with  plates  of 
pure  gold  and  silver,  which  accounts  for  the  metals 
not  being  as  soluble  as  in  ores.  This  is  accounted  for 
by  the  unavoidable  polarization  of  the  gold  and  silver 
plates  by  hydrogen  bubbles  that  are  formed  during 
the  experiment.  The  curve,  Fig.  4,  clearly  shows 
the  comparative  solubility  of  the  different  metals 
and  minerals  in  cyanide  solutions  of  different 
strengths  and  the  efficiency  of  KCN  per  unit  in  each 
case.  On  studying  the  table  (Fig.  4)  it  will  be 
seen  that  no  proportionality  exists  between  the 
weight  of  metal  dissolved  and  the  percentage 
strength  of  solution.  Thus  a  0.01%  solution  dis- 
solves 24  parts  of  gold,  whereas  a  0.1%  solution, 
ten  times  as  strong,  dissolves  only  82.5  parts  in  that 
time.  Therefore,  the  weaker  the  solution  the  greater 
is  the  efficiency  per  molecule  of  KCN  dissolved,  but 


V 


?*9*u9ar   #3^ 


CYANIDATION 


31 


in  practical  work  this  rule  can  easily  be  carried  too 
far.  A  solution  that  would  have  the  maximum  dis- 
solving power  and  a  maximum  efficiency  per  molecule 
of  KCN  is  desired  for  working  strength.  This  can, 
however,  be  arrived  at  only  in  an  empirical  way  as 
the  time  factor  has  to  be  taken  into  account.* 

TABLE  1. 


Order  of 
Solubilities 

Per  Cent  KCN 

0.1 

0.09 

0.08 

0.07 

0.06 

0.05 

0.04 

0.03 

0.02 

0.01 

Gold 
Silver 

82.5 
45.5 

78.5 
43.3 

74.0 
40.8 

60.25 
38.2 

64.0 
35.3 

58.0 
32.0 

51.5 
28.4 

44.5 
24.6 

35.5 
19.5 

24.0 
13.2 

Efficiency 
of  KCN  per 
unit 

0.344 

0.363 

0.385 

0.412 

0.444 

0.483 

0.536 

0.618 

0.740 

1.000 

Now  if  the  time  required  to  treat  an  ore  with  solu- 
tions of  known  strength  be  known,  and  it  be  desired 
to  vary  the  capacity  of  the  works  by  lengthening  or 
shortening  the  time  of  treatment,  what  strength  of 
solution  would  give  equally  good  results  may  then 
be  determined  from  the  data  given.  Thus  if  it  re- 
quired 200  hours  to  treat  an  ore  effectively  with  a 
0.01%  solution,  and  it  became  necessary  to  increase 
the  capacity  of  the  works  by  shortening  the  time  of 
treatment  to  70  hours  without  decreasing  the  extrac- 
tion, it  would  be  necessary  to  find  in  the  column  a 
dissolving  efficiency  unit  which  multiplied  by  200 
equals  70  or  thereabouts.  Then  the  strength  of  solu- 

*If  the  time  factor  is  left  out  of  the  question,  as  in  cases 
where  the  gold  and  silver  dissolve  rapidly  but  can  be 
washed  out  of  the  ore  only  slowly,  the  most  efficient  strength 
of  solution  is  then  from  0.07  to  0.09%  KCN. 


32  TESTING  FOR  PROCESSES 

tion  at  the  head  of  the  column  is  the  one  required. 
In  this  case  it  will  be  found  that  0.344  X  200  =  68.8. 
The  required  strength  found  at  the  head  of  the  col- 
umn is  0.1  per  cent. 

4.     FINENESS  OF  GRINDING 

Take  3  samples  of  the  ore  ground  so  as  to  pass 
through  20,  40,  and  100  mesh,  respectively;  place  in 
bottles  with  the  required  amount  of  alkali  and  cya- 
nide solution  as  determined  in  the  previous  tests; 
agitate  for  48  hours ;  filter ;  wash  ;  dry ;  assay  the  tail- 
ing. Calculate  percentage  of  extraction  for  each  mesh. 

The  object  is  to  break  the  ore  into  particles  of  such 
size  as  will  leave  the  gold  and  silver  capable  of  com- 
ing into  contact  with  the  cyanide  solution.  This  end 
is  obtained  by  fine  pulverizing,  and  much  may  be 
learned  by  examining  the  grains  under  a  low-power 
microscope  to  determine  the  minimum  degree  of  fine- 
ness necessary;  a  point  which,  however,  must  ulti- 
mately be  decided  from  empirical  results  by  testing 
with  cyanide  solution.  The  more  finely  the  ore  is 
crushed,  the  more  perfectly  does  the  gold  and  silver 
content  dissolve,  but  if  crushed  too  fine  an  excess 
of  dust  or  slime  forms,  which  interferes  with  the  sep- 
aration of  the  dissolved  gold  and  silver  and  causes 
an  increased  consumption  of  cyanide.  If  crushed  too 
coarse  a  large  percentage  of  gold  and  silver  remains 
locked  up  and  is  not  dissolved,  but  that  portion 
which  does  dissolve  may  be  separated  more  or  less 
readily.  Under  no  circumstances  can  an  ore  be 
crushed  to  one  uniform  size,  but  the  aim  is  to  crush 
it  so  as  to  get  a  maximum  number  of  those  particles 


CYANIDATION  33 

that  give  the  highest  extraction  at  the  lowest  cost. 
This  can  be  found  only  by  experiments.  In  many 
cases  it  is  advantageous  to  crush  wet,  while  in  others 
dry  crushing  may  be  more  beneficial  or  profitable. 

5.     TIME  OF  TREATMENT 

Take  4  samples  ground  through  the  mesh  as  de- 
termined in  (4)  ;  add  alkali  and  cyanide  solution  as 
given  under  (2).  Agitate  occasionally  for  1,  3,  5, 
and  10  days,  respectively,  with  the  free  access  of  air. 
Determine  the  percentage  of  extraction  for  each 
sample.  Construct  a  curve  as  shown  in  Fig.  5  to 
show  the  rate  of  dissolution  of  the  gold  and  silver. 
The  curve  will  help  to  indicate  when  the  profitable 
time  limit  of  treatment  would  be  reached. 

In  an  ore  the  variety  of  sizes  and  shapes  of  the 
particles  of  gold  must  be  very  great,  but  a  consid- 
eration of  a  few  assumed  cases  will  assist  in  under- 
standing what  actually  takes  place  in  practical 
operations.  (1)  Where  the  gold  is  in  thin  plates 
the  rate  of  dissolution  will  be  uniform  until  complete. 
Where  gold  is  imbedded  in  pyrites  so  that  only  one 
edge  can  be  attacked,  it  would  also  have  a  constant 
rate  of  dissolution.  (2)  In  the  other  extreme  case 
where  the  gold  is  assumed  to  exist  in  spheres,  then, 
in  equal  intervals  of  time,  equal  thickness  of  shell 
will  be  dissolved  but  as  each  successive  shell  is  of 
smaller  diameter  the  amount  of  gold  dissolved  in 
an  equal  space  of  time  will  be  continually  decreased. 
(3)  If,  however,  in  order  to  approximate  more  closely 
to  actual  conditions,  a  mixture  of  spheres  of  various 
sizes  be  assumed,  it  is  clear  that  the  smaller  spheres 


U 


\ 


\ 


\ 


§1. 


CYANIDATION  35 

will  be  entirely  dissolved  before  the  larger  ones  and 
in  such  cases  the  rate  at  the  beginning  will  be  con- 
siderably increased. 

Percolation  and  Leaching. — The  process  of  causing 
a  liquid  to  pass  through  a  porous  medium  such  as 
sand  is  termed  percolation,  and  when  the  liquid,  as 
it  percolates,  dissolves  some  of  the  porous  medium, 
it  is  said  to  leach.  The  advantages  of  the  percolation 
method  of  leaching  in  tanks  are,  that  enormous 
quantities  may  be  dealt  with  in  one  operation,  and 
by  the  simplest  possible  plant  at  the  lowest  cost; 
also  the  method  admits  of  easy  and  close  control. 

6.     RATE  OF  PERCOLATION 

This  varies  with  the  volume  and  uniformity  of 
the  interstices  between  the  ore  particles,  and  to  some 
extent  with  the  depth  of  the  column,  the  pressure, 
and  temperature.  The  constituents  of  the  ore  also 
have  some  influence  on  the  rate  of  percolation,  for 
some  minerals  such  as  galena,  pyrite,  and  blende 
have  a  marked  effect  in  diminishing  the  rate  of  per- 
colation, and  this  therefore  varies  with  the  propor- 
tion of  these  minerals  to  the  silicious  particles  pres- 
ent. It  naturally  follows  that  the  coarser  the  ore 
particles  are,  the  larger  will  be  the  pores  or  inter- 
stices and  the  more  rapid  the  rate  of  percolation. 
Then,  again,  uniformity  in  the  size  of  the  ore  par- 
ticles has  something  to  do  with  the  rate  of  percola- 
tion and  the  leaching  effect,  for  if  an  ore  be  crushed 
so  as  to  pass  through  a  90-mesh  screen,  and  washed 
free  of  slime,  a  good  leachable  product  may  be  ob- 
tained. If  the  fine  ore  be  mixed  with  coarser  par- 


36  TESTING  FOR  PROCESSES 

tides,  percolation  becomes  slower  than  when  the  two 
sizes  are  treated  separately,  and  leaching  is  retarded. 

The  loss  in  velocity  as  the  solution  descends  into 
the  ore,  seems  to  be  due  in  a  large  measure  to  the 
pushing  forward  of  air  contained  between  the  par- 
ticles, but  also  the  skin  friction  between  the  solution 
and  the  ore.  In  its  attempt  to  escape,  some  of  the 
air  moves  in  an  upward  direction  and  sometimes 
reaches  the  surface,  but  a  very  large  number  of  the 
globules  are  retained,  as  the  pressure  does  not  enable 
them  to  force  their  way  upward  and  escape.  These 
remain  in  the  ore,  balanced  with  the  weight  of  solu- 
tion above.  Such  a  state  of  things  has  a  marked 
detrimental  effect  on  the  rate  of  leaching  and  causes 
an  excess  of  moisture  to  be  left  in  the  ore  after 
draining.  With  upward  percolation  these  globules 
are  not  formed  to  the  same  extent  as  in  the  down- 
ward method.  The  chief  quantity  of  air  below  the 
ore  surface  escapes  through  the  outlet  pipe  below 
the  filter,  and  if  this  outlet  pipe  is  not  open  a  large 
volume  of  air  bubbles  upward,  through  the  sands, 
affecting  the  texture  of  the  ore. 

The  amount  of  moisture  retained  by  a  leached  ore 
depends  on  the  size  of  the  particles,  the  uniformity 
of  texture,  air  pressure,  and  temperature,  and  to 
some  extent  on  the  depth  of  the  ore  column.  Tem- 
perature has  an  effect,  which  in  some  cases  is  marked. 
The  higher  the  temperature,  the  smaller  is  the  quan- 
tity of  moisture  retained.  Bad  leaching  is  chiefly 
the  result  of  uneven  percolation,  and  uneven  perco- 
lation follows  from  want  of  uniformity  in  the  texture 
of  the  ore,  or  from  want  of  uniformity  in  the  pores 


CYANIDATION  37 

of  the  filter  cloth.  Whenever  the  filter  cloth  becomes 
clogged  with  slime  the  extraction  becomes  very  poor. 

There  is  no  fixed  rule  for  applying  solution  to 
the  ore.  Some  operators  apply  the  solution  in  three 
or  four  large  charges,  while  others  apply  it  in  many 
small  charges.  Theoretically  the  latter  method 
should  give  the  better  extraction,  but  in  practice  the 
former  gives  equally  good  results.  It  is  in  the  final 
treatment,  when  the  water  wash  is  applied,  that  a 
number  of  small  charges  are  highly  beneficial.  Three 
or  four  small  charges  of  10  tons  each  will  extract 
more  salt  absorbed  by  the  ore  than  one  large  charge 
of  30  to  40  tons.  This  is  noticeable  in  the  first  small 
charge  of  water  wash,  which  after  percolating  will 
be  found,  when  drawn  off,  to  be  richer  in  alkali, 
cyanide,  and  gold  than  the  previous  weak  solution. 
The  cause  of  this  is  not  clear,  but  it  would  appear 
to  be  due  to  some  reaction  between  the  absorbed  salt 
and  the  air  drawn  into  the  ore  mass  as  the  solution 
drains  off,  whereby  the  absorbed  salt  becomes  more 
soluble. 

If  the  ore  is  to  be  leached  by  percolation  the  solu- 
tion must  sink  through  the  column  of  ore  with  suffi- 
cient rapidity  to  keep  the  time  of  treatment  of  one 
lot  of  ore  within  commercial  limits.  The  rate  of 
percolation  is  determined  sometimes  by  the  time  the 
solution  above  the  ore  sinks  to  a  certain  depth ;  thus 
over  3  inches  per  hour  is  good,  1%  inches  is  fair,  and 
%  of  an  inch  is  poor.  Less  than  %  of  an  inch  is 
usually  uneconomical.  A  better  method  is  to  judge 
the  porosity  of  the  ore  by  comparing  the  quantity 
of  moisture  retained  after  draining.  The  greater 


38  TESTING  FOR  PROCESSES 

the  percentage  of  moisture  retained  the  smaller  the 
porosity  and  the  slower  the  percolation. 

Since  the  cost  of  leaching  in  tanks  is  less  than  by 
any  other  means,  it  is  therefore  advisable  to  treat 
the  largest  quantity  practicable  by  this  method.  In 
some  classes  of  ore  the  quantity  of  slime  is  of  such 
a  character  and  so  small  after  reducing  the  ore  to 
the  required  fineness  that,  when  mixed  evenly  with 
the  leaching  products,  it  does  not  materially  inter- 
fere with  the  percolation,  or  with  the  extraction  of 
the  metal.  In  such  cases  it  might  be  most  economical 
not  to  size  the  ore  before  treatment.  In  case  it  is 
found  necessary  to  size  the  ore  to  secure  the  necessary 
rate  of  percolation,  the  maximum  quantity  amen- 
able to  this  method  of  treatment  is  determined  by 
first  sizing  a  sample  and  testing  each  size  separately, 
when  the  limit  will  be  found  at  which  percolation 
in  tanks  must  cease. 

Test. — To  compare  the  porosity  and  rate  of  per- 
colation, the  apparatus  shown  in  Fig.  6  may  be  used. 
(A)  is  a  tube  from  1  to  2  inches  in  diameter  and 
4  or  5  ft.  long,  connected  at  its  lower  end  to  a 
measuring  glass  (B)  by  means  of  rubber  tubing  (C). 
The  tube  (A)  is  filled  to  a  mark,  say  4  ft.  from  the 
bottom,  with  ore  previously  dried  and  the  net  weight 
noted.  A  rubber  stopper  (a)  fitted  with  a  short 
length  of  glass  tube  serves  as  a  means  of  connection 
(c).  A  piece  of  muslin  is  placed  over  the  stopper 
(a)  to  support  the  sand  and  act  as  filter-bottom. 
The  glass  (B)  and  the  tube  (c)  are  filled  with  clean 
cyanide  solution  of  work-strength  and  zero  mark  is 
brought  to  the  level  (a).  The  cock  (T)  is  opened 


CYANIDATION 


39 


and  (B)  is  very  gradually  raised  until  the  solution 
in  (A)  rises  to  a  point  about  three  inches  above  the 
ore  surface  and  the  solution  in  (B)  is  on  the  same 


Fig.  6.  APPARATUS  FOR  TESTING  THE  POROSITY  OF 
LEACHING  MATERIAL 

level.  The  amount  of  solution  left  in  (B)  is  read, 
thus  giving  the  volume  of  solution  necessary  to  cover 
an  amount  of  ore  equal  to  the  volume  contained  in 
the  tube  (A).  (B)  and  (c)  are  next  lowered  and 
the  solution  allowed  to  drain  off.  Time  observations 


40  TfeSTING  FOR  PROCESSES 

are  then  made  at  the  rate  of  fall  of  the  solution. 
Allow  the  ore  to  drain  for  as  many  minutes  per  foot 
of  depth  as  would  be  employed  on  a  large  scale. 
Finally  disconnect  (C)  from  (a)  and  weigh  (A)  to 
find  the  weight  of  moisture  retained  in  the  ore. 

7.     SIZING  TEST 

In  case  the  rate  of  percolation  as  found  above  does 
not  come  within  the  limit  set  for  the  ore,  it  will  gen- 
erally be  advantageous  to  classify  the  ore  into  two 
or  more  sizes,  coarse  or  fine.  Slime  may  be  defined 
as  that  ore  which  cannot  be  economically  treated  in 
the  leaching  tanks  on  account  of  the  fineness  of  the 
particles.  The  next  larger  size  resulting  from  the 
classification  should  not  hold  more  than  40%  moist- 
ure. 

The  process  of  wet-sizing  is  best  carried  out  in 
some  form  of  elutriation  apparatus,  but  it  may  be 
performed  in  a  less  perfect  manner  by  the  use  of  a 
gold  pan.  In  case  the  latter  method  is  used,  the  ore 
is  placed  in  a  gold  pan  and  the  slime  floated  off 
from  the  sand  in  very  much  the  same  manner  as 
given  on  page  12.  A  much  more  satisfactory  result 
is  obtained  by  employing  a  small  elutriation  appara- 
tus, as  shown  in  Fig.  7.  The  apparatus  consists  of 
inverted  cones  having  an  angle  of  about  33°  at  their 
apexes,  and  having  diameters  of  bases  in  the  pro- 
portion of  1 :  2 :  4.  Other  proportions  are  not  of 
much  importance.  The  vessel  (A)  is  adjusted  to  a 
suitable  height  and  water  is  allowed  to  flow  into  it 
a  little  faster  than  it  flows  out  through  the  tube  (a), 
and  is  thus  kept  overflowing.  When  the  apparatus 


CYANIDATION 


41 


is  filled  with  water,  a  clip  (s)  is  closed  and  a  small 
quantity  of  ore  in  the  form  of  sludge  is  fed  from 
the  funnel  into  the  first  cone  by  loosening  the  spring 
clip  (p).  The  clip  (p)  is  then  closed  and  (s)  is 
opened  and  the  water  allowed  to  flow  from  the  cis- 
tern (A)  until  it  is  practically  clear  as  it  leaves  the 
discharge  pipe  (d).  At  the  apex  of  each  cone  is  a 
short  tube  of  rubber  fitted  with  a  spring  clip  for 


Fig.  7.     ELUTRIATION  APPARATUS  FOR  SIZING  AND 
CLASSIFYING 

the  purpose  of  discharging  the  ore  into  separate  ves- 
sels at  intervals.  The  operation  is  repeated  until  a 
sufficient  quantity  of  each  size  is  collected.  The  fine 
and  light  particles  flow  into  the  cistern  (D)  where 
they  are  allowed  to  settle.  By  re-treating  the  settled 
portion  in  (D),  seven  sizes  may  be  obtained;  and  if 
each  size  is  submitted  to  a  percolation  test,  the  pro- 
portion of  non-leachable  material  present  and  in 
what  sizes  the  gold  and  silver  chiefly  exist  may  be 
readily  determined.  Other  factors  may  also  be  ren- 


42  TESTING  FOR  PROCESSES 

dered  apparent,  as  for  instance,  whether  the  ore  has 
been  crushed  sufficiently  fine  to  liberate  the  gold 
and  silver.  Tests  may  be  made  on  each  size  by  leach- 
ing with  cyanide  to  determine  where  the  bulk  of 
metal  remaining  in  the  residues  is  to  be  found. 

8.     SLIMING  TEST 

In  case  the  solution  of  the  gold  and  silver  is  not 
obtained  in  the  coarser  sizes  of  the  ore,  tests  should 
be  made  to  determine  the  effect  of  sliming  all  or 
part  of  the  ore  as  may  be  found  necessary  to  obtain 
a  good  extraction. 

Test. — Grind  a  300-gram  sample  of  the  ore  so  as  to 
pass  through  a  200-mesh  screen,  introduce  into  an  agi- 
tation apparatus,  as  shown  in  Fig.  8.  Add  the  proper 
amount  of  alkali  and  fill  the  apparatus  to  within  3 
inches  of  the  top  with  a  measured  volume  of  0.01  to 
0.05%  strength  cyanide  solution.  Turn  on  the  air 
gently  so  as  to  make  an  effective  circulation  of  the 
slime  but  not  so  violent  as  to  cause  loss  by  spattering. 
In  treating  gold  ore  slime  in  the  laboratory,  agitation 
from  2  to  4  hours  is  generally  sufficient.  With  silver 
ores,  the  time  required  may  be  several  weeks.  Dur- 
ing the  agitation  take  a  sample  by  means  of  a  siphon, 
(one  shown  in  Fig.  8  is  very  convenient)  every  hour; 
assay  for  gold,  silver,  and  free  KCN.  Return  the 
unused  portion  of  each  sample  to  the  agitation  ap- 
paratus, also  note  the  amount  used  for  each  assay. 
Keep  the  strength  of  the  KCN  solution  up  to  the 
standard  selected  by  the  addition  of  known  quan- 
tities of  KCN.  From  the  assays  of  the  samples  taken 
above,  determine  the  percentage  of  the  extraction 


CYANIDATION 


43 


for  each  hour  and  construct  a  curve  from  these  data 
as  shown  in  Fig.  5,  p.  34.     The  amount  of  cyanide 


Fig.  8.     SLIME  AGITATOR 

added    to   keep   the   solution   up   to   the   standard 
strength  will  give  an  approximate  idea  of  the  amount 


44 


TESTING  FOR  PROCESSES 


of  cyanide  consumed  by  the  weight  of  the  slime 
treated.  Calculate  in  terms  of  KCN  per  ton  of  dry 
slime.  See  footnote,  page  27. 

9.     DECANTATION  TESTS 

Weigh  out  five  500-gram  samples  of  the  slime,  put 
each  into  a  1000  c.c.  measuring  cylinder  together  with 
different  percentages  of  some  coagulating  substance, 
such  as  lime,  say  0.1,  0.2,  0.4,  0.6,  0.8%,  add  dilute 
cyanide  solution,  say  0.05%,  a  little  at  first,  and  thor- 


Fig.  9.     SAMPLING  SIPHON 

oughly  agitate,  then  fill  up  to  the  1000  c.c.  mark.  Al- 
low these  to  stand  3  or  4  hours,  taking  care  that  the 
temperature  remain  constant,  and  note  the  depth  of 
sediment  in  each  case.  The  best  percentage  of  lime  is 
that  which  gives  the  greatest  amount  of  clear  solu- 
tion for  decantation.  In  some  cases  a  maximum 


CYANIDATION  45 

point  is  soon  reached,  while  in  others  the  most  eco- 
nomical amount  of  lime  has  to  be  determined  partly 
by  cost.  Beyond  a  certain  amount  of  lime  no  ad- 
vantage is  derived,  but  often  subsidence  is  retarded. 
Caustic  lime  is  generally  employed  in  the  cyanide 
process,  as  it  protects  the  cyanide,  but  it  is  not  so 
good  a  coagulating  substance  as  other  compounds 
of  calcium.  Thus  calcium  carbonate  (chalk)  is  often 
more  rapid  and  produces  a  more  dense  sediment, 
which  is  not  so  easily  disturbed  while  decanting  as 
when  hydrate  is  used. 

10.     DETERMINATION  OF  THE  CAUSES  OF  CYANIDE 
CONSUMPTION 

The  cause  of  a  high  cyanide  consumption  may  be 
determined  by  an  analysis  of  the  solution  after  use. 
For  every  part  of  cyanide  rendered  inoperative  a 
corresponding  part  of  metal  enters  solution.  Thus 
one  part  by  weight  of  iron  consumes  seven  parts  by 
weight  of  cyanide,  and  zinc  and  copper,  the  number 
of  parts  shown  in  the  curve,  Fig.  4,  page  30.  Salts 
of  aluminum  and  magnesium  form  hydrates  with  a 
liberation  of  HCN  as  below : 

MgS04+2KCN-f2H20=Mg(OH)2+K2S04+2HCN. 

This  may  be   overcome  by  a   preliminary  alkaline 

treatment  during  the  crushing,  their  hydrates  being 

precipitated  which  are  inert  toward  KCN  as  follows : 

MgS04  +  Ca(OH)2  =  Mg(OH)2  +  CaS04. 

Any  free  acid  has  a  powerful  effect : 

2KCN  +  H2S04  =  2HCN  +  K2SO4. 

Cocoa  matting,  jute,  and  vasculose  when  new  arc 


46  TESTING  FOR  PROCESSES 

especially  destructive  to  the  solution.  Vasculose  acts 
as  a  reducing  agent.  The  presence  of  soluble  and 
insoluble  carbonates  greatly  facilitate  the  decompo- 
sition of  the  cyanides.  Metallic  iron  from  the  crush- 
ing machinery  becomes  disseminated  throughout  the 
ore  in  fine  grains.  It  dissolves  slowly  but  is  never- 
theless a  factor  in  the  decomposition  of  cyanide,  and 
the  hydrogen  thus  liberated  combines  with  the  valu- 
able oxygen  of  the  solution  as  follows : 

Fe  +  6KCN  +  2H2O  =  KFe(CN)6  +  2KOH  +  2H. 
.    2H  +  0  =  H20. 

Pyrite  oxidizes  rapidly  in  moist  air  to  ferrous  sul- 
phate which  is  the  cause  of  considerable  consump- 
tion of  cyanide.  Most  of  this  injurious  substance 
may  be  removed  by  a  preliminary  water-wash  and 
the  remainder  neutralized  by  alkali,  forming  ferrous 
hydrate  and  alkaline  sulphate.  Ferrous  hydrate  and 
the  sulphate  act  as  de-oxidizers  and  consume  oxygen 
that  might  otherwise  be  of  service  in  dissolving  the 
gold  and  silver.  Ferrous  sulphate  oxidizes  to  nor- 
mal and  basic  ferric  sulphates,  which  in  time  con- 
sume cyanide,  forming  K4Fe(CN)6  and  probably 
K3Fe(CN)8. 

Copper  in  the  native  state  and  in  the  form  of  dyna- 
mite caps  is  also  acted  upon  by  cyanide  as  below : 
2Cu  +  4KCN  +  2H2O  =  K2Cu2(CN)4+  2KOH  +  2H. 
2H  +  0  =  H20. 

A  mere  chemical  analysis  is  not  to  be  relied  upon 
in  testing  copper  ores,  as  much  depends  upon  the 
physical  condition  in  which  the  copper  and  gold 
exist.  Sometimes  an  ore  containing  from  2  to  3% 


CYANIDATION  47 

of  copper  may  be  successfully  treated  by  cyanide, 
while  some  ores  containing  0.05%  and  less  are  fail- 
ures. 

Arsenic  is  not  harmful  to  the  cyanide  process,  but 
nearly  all  of  its  compounds  are  soluble  in  caustic 
alkali  and  decompose  cyanide  in  the  absence  of  free 
alkali.  Mispickel  (arsenical  pyrite,  FeAsFeS3)  is 
next  to  iron  pyrite  the  most  abundant  metallic  com- 
pound found  associated  with  gold  ores,  and  is  gen- 
erally the  most  auriferous  of  the  two.  It  weathers 
in  moist  atmospheres  to  ferrous  sulphate  and  to 
hydrate  and  oxide  of  arsenic.  The  latter  is  only 
harmful  in  the  absence  of  free  alkali.  In  treating 
mispickel  ores,  it  is  generally  found  that  the  addition 
of  large  amounts  of  lime  to  the  ore  effects  a  large 
saving  of  cyanide  and  improves  the  extraction.  If 
magnesia  is  procurable  it  will  be  found  to  answer 
equally  as  well,  and,  as  a  rule,  a  better  extraction 
can  be  relied  upon.  Sulphides  of  arsenic,  as  realgar 
and  orpiment,  are  both  attacked  by  the  alkalies, 
forming  arsenites  and  thio-arsenites.  Alloys  of  iron 
and  arsenic  must  be  finely  ground  for  a  good  extrac- 
tion of  the  gold  and  silver.  The  compounds  of  alkali 
and  arsenic  that  form  act  as  reducers,  and  consume 
oxygen  dissolved  in  the  cyanide  solution,  and  in  this 
way  retard  the  dissolution  of  gold  and  silver. 

Antimony  sulphide,  as  stibnite,  is  often  found  as- 
sociated with  gold  ores  and 'is  often  highly  aurifer- 
ous. Like  arsenic,  antimony  does  not  form  any 
definite  compound  with  cyanide,  but  the  sulphide  is 
very  soluble  in  caustic  alkali  and  decomposes  cya- 
nide. The  antimony  compounds  act  as  strong  de- 


48  TESTING  FOR  PROCESSES 

oxidizers  and  remove  the  absorbed  oxygen  from  the 
solution  by  forming  antimonate  and  thio-antimonate. 
Gold  and  silver  may  be  rendered  almost  insoluble 
from  this  cause ;  also  from  the  fact  that  much  cyanide 
is  destroyed.  Auriferous  antimonial  concentrate 
carrying  as  high  as  20%  antimony,  has  been  suc- 
cessfully treated  by  giving  the  ore  an  ordinary  care- 
ful roast,  previously  mixed  with  2  to  5%  of  charcoal 
or  c,oal,  followed  by  a  hot  acid  wash  of  dilute  hydro- 
chloric acid  (obtained  by  exposing  old  chlorine  solu- 
tions to  the  direct  sunlight)  and  then  lixiviating 
with  cyanide  solution.  Another  method  is  to  roast 
as  above,  salt  at  the  end  of  the  roast,  and  chlorinate. 
Both  methods  have  yielded  90%  extraction  but  have 
the  drawback  of  requiring  very  careful  roasting. 

Tellurium  is  often  found  in  minerals  associated 
with  gold  ores,  which  are  usually  highly  auriferous 
or  argentiferous.  The  presence  of  tellurides,  when 
they  form  any  appreciable  proportion  of  the  metallic 
minerals,  usually  makes  an  ore  difficult  to  treat,  as 
the  action  of  the  cyanide  on  the  gold  is  slow.  The 
cause  of  this  slowness  of  action  is  not  altogether 
apparent,  but  it  is  evident  that  there  is  a  wide  differ- 
ence in  the  solubility  of  the  gold  in  different  telluride 
minerals  and  even  in  the  same  minerals  from  differ- 
ent localities.  There  is  no  action  between  the  cyan- 
ogen radical  and  the  tellurium,  but  the  alkali  in 
presence  of  oxygen  invariably  dissolves  some  of  the 
metal  and  also  acts  on  the  sulphur  when  in  combina- 
tion, forming  a  solution  which  has  a  reducing  action. 
When  a  telluride  ore  is  roasted,  it  leaves  a  residue 
containing  Te02,  and  this  oxide  is  readily  soluble  in 


CYANIDATION  49 

KOH,  forming  a  tellurite,  which  also  acts  as  a  re- 
ducing agent,  and  absorbs  oxygen  from  the  solution. 
The  same  change  takes  place  with  KCN,  with  the  evo- 
lution of  HCN.  Roasted  tellurides  are,  however, 
capable  of  being  treated  and  the  gold  extracted  with 
good  results. 

In  cyaniding  tailing  from  amalgamation  mills,  me- 
tallic mercury  finds  its  way  into  the  cyanide  vats 
along  with  the  tailing  and  is  acted  upon  by  the  cya- 
nide. Mercury  has,  however,  a  greater  affinity  for 
sulphur  than  for  cyanogen,  and  as  the  solutions  usu- 
ally contain  sulphides  and  sulphates,  the  presence  of 
mercury  is  not  altogether  undesirable. 

Zinc  blende  and  galena  have  little  effect  on  the 
cyanide  solution. 

Quantitative  Analysis  of  the  Cyanide  Solution. — Place 
50  grams  of  the  ore  in  a  wide-mouthed  bottle,  add 
100  c.c.  of  KCN  solution,  agitate  for  15  hours ;  filter ; 
take  20  c.c.  of  solution,  equivalent  to  10  grams  of  ore, 
and  determine  its  constituents.  The  metals  in  solu- 
tion should  be  determined  quantitatively  by  the  usual 
methods,  but  for  ordinary  purposes  a  qualitative 
analysis  is  sufficient.  Since  alkaline  suphides  act  so 
injuriously  on  the  cyanide  solution  during  the  leach- 
ing, it  is  important  to  detect  their  presence.  They 
are  soluble  in  water.  The  most  delicate  test  is  by 
means  of  the  nitro-prussides.  They  are  formed  by 
adding  nitric  acid  to  a  solution  of  ferro  or  ferri- 
cyanide  of  potassium.  Add  a  few  drops  of  the  nitro- 
prusside  to  the  cyanide  solution.  If  any  alkaline 
sulphide  is  present,  even  in  the  minutest  quantity, 
the  solution  will  assume  a  brilliant  purple  color. 


50  TESTING  FOR  PROCESSES 

GENERAL  REMARKS 

CAUSES   FOR  THE   NON-EXTRACTION 

(1).  Gold  alloyed  or  in  combination. — These  com- 
binations may  be  broken  up  by  a  preliminary  roast, 
or  extraction  effected  by  fine  grinding,  long  contact 
with  the  cyanide  solution,  or  both. 

(2).  Large  consumption  of  KCN  (see  'Determina- 
tion of  the  Causes  of  Cyanide  Consumption',  page 
45). 

(3).  Coarse  gold. — This  may  be  amalgamated  or 
concentrated  before  cyaniding. 

(4).  Presence  of  soluble  sulphides. — Add  a  solu- 
ble lead  salt  or  an  oxidizing  agent. 

(5).  If  the  ore  contains  much  kaolin  or  talc. — If 
coarse  crushing  is  ineffectual,  the  separation  of  the 
sand  and  slime  does  not  prove  effectual,  or  the  slime 
cannot  be  filtered  or  decanted,  the  ore  is  not  adapted 
to  economical  cyanidation. 

(6).  Ores  containing  considerable  oxidized  copper 
are  not  adapted  to  the  cyanide  process,  without  a 
preliminary  treatment  as  an  acid  wash.  The  addition 
of  ammonium  salts  as  NH4C1  is  usually  beneficial  in 
treating  ore  containing  copper.  The  ammonia- 
cyanide  process  may  be  used  in  treating  copper, 
nickel,  and  zinc  ores  or  tailings  containing  values  of 
gold  and  silver.  Ammonia  is  used  as  a  solvent  for 
the  base  metals  in  the  ore.  Potassium  cyanide  is 
added  to  the  ammonia,  forming  in  the  presence  of 
copper,  cupric  cyanide  (Cu(CN)2)  and  an  excess  of 
soluble  copper  hydroxide  (Cu(OH)2),  which  com- 
bines the  extraction  of  the  gold  and  silver  with  the 


CYANIDATION  51 

copper.  The  ammonia-cyanide  solution  is  prepared 
by  dissolving  1  to  4  pounds  of  KCN  in  a  ton  of  solu- 
tion \%  of  NH3  for  each  per  cent  of  copper  in  the 
ore.  With  the  use  of  open  leaching  tanks,  a  solution 
containing  more  than  1%  NH3  is  not  to  be  recom- 
mended on  account  of  loss  by  volatilization,  hence 
the  excess  copper  should  be  removed  by  preliminary 
leaching  with  acid  or  water.  Ammonia  cupric  cya- 
nide is  very  stable,  has  a  greater  dissolving  efficiency 
than  ordinary  cyanide,  and  has  important  oxidizing 
powers.  This  is  a  desirable  quality  in  cyaniding, 
especially  in  the  presence  of  ferrous  and  other  re- 
ducing salts. 

(7).  Ores  containing  heavy  minerals  injurious  to 
the  cyanide  solution. — Concentration  before  cyanida- 
tion  may  make  tailing  amenable  to  the  process. 

CYANIDE  POISONING 

For  an  antidote,  drink  ferrous  sulphate  followed 
by  sodium  bicarbonate.  Next  take  a  purgative. 
Walk  about  and  prevent  sleep.  KCN  in  fresh  cuts 
is  liable  to  form  painful  sores. 

REACTIONS  IN  THE  CYANIDE  PROCESS 
IN  THE  PRELIMINARY  TREATMENT 

Fe203S03  +  GNaOH  =  2Fe(OH)3  +  3Na2S04 
FeS04  +  Ca(OH)2  =  Fe(OH)2  +  CaS04 
ZnSO4  +  2NaOH  =  Zn(OH)2  +  CaSO4 

IN  THE  SOLUTION  TANKS 

4KCN  +  2Au  +  H20  +  0  =  2KAu(CN)2  +  2KOH 
(Ellsner's  Equation) 


52  TESTING  FOR  PROCESSES 

4KCN  +  2Ag  +  H20  +  0  =  2KCN.2AgCN  +  2KOH 

(Ellsner's  Equation) 
2KCN  +  FeS04  =  Fe(CN)2  +  K2S04 
Fe(CN)2  +  4KCN==K4Fe(CN)6 
Pe2(S04)8  +  6KCN  +  6H20  =  2Fe(OH)3  +  3K2S04 

+  6HCN 
A12(S04)3  +  3H20  +  6KCN  =  A1203  +  3K2S04  + 

6HCN 

IN  THE  AIR 

2KCN  +  C02  +  H20  =  2HCN  +  K2C03 
KCN  +  0  =  KCNO 
2KCNO  +  30  =  K2C03  +  C02  +  2N 
2KCN  +  2H20  =  2HCN  +  2KOH 

SUBSTANCES  IN  SOLUTION 

Zn(CN)2  +  4KOH  (excess)  =  K2O.ZnO  +  2KCN  + 
2H20 

K2Zn(CN)4  +  Ca(OH)2=2KCN  +  Ca(CN)2Zn(OH)2 

IN  PRECIPITATING  TANKS  AND  ZINC  BOXES 
2KAu(CN)2  +  Zn  =  K2Zn(CN)4  +  2Au 
2H20  +  Zn  =  Zn  (OH)  2  +  2H 
Zn(OH)2  +  2KCN  =  Zn(CN)2  +  2KOH 
Zn(CN)2  +  2KCN  =  K2Zn(CN)4 
2KOH  +  CO2  =  K2C03  +  H20 
HCN  +  KOH  =  KCN  +  H20 

Electrolysis  will  give  from  KAu(CN)2  =  Au  +  K 
+  2CN, 


2KOH  +  Zn  =  ZnK202  +  2H. 
Report.  —  Make  out  a  final  report  of  all  tests  made 
in  a  neat  form  with  all  notes  and  observations  that 


CYAN  I  DAT  I  ON  53 

may  be  of  value  in  the  further  treatment  of  the  ore. 
Compare  with  the  net  returns  that  would  result  from 
the  use  of  other  available  methods  such  as  amalga- 
mation, chlorination,  shipping  to  smelter  direct,  con- 
centration and  shipping  concentrates  to  the  smelter. 

CYANIDE  TEST  ON  LOTS  OF  30  POUNDS  AND 
MORE  OF  ORE 

The  procedure  in  this  case  is  conditioned  upon 
the  results  obtained  in  the  preliminary  tests.  A 
scheme  of  treatment  for  each  ore  must  be  devised. 
If  the  ore  is  to  be  treated  by  percolation  the  follow- 
ing suggestions  will  be  of  value. 

Take  30  pounds  of  ore  and  crush  to  the  required 
mesh,  being  careful  to  add  the  proper  amount  of 
alkali.  Spread  the  ore  evenly  over  the  bottom  of 
one  of  the  tanks  in  the  leaching  plant,  shown  in  Fig. 
10.  Ten  pounds  of  strong  solution  (0.25%  KCN)  is 
run  on  and  is  immediately  drawn  off  so  that  the 
solution  just  covers  the  ore.  It  is  allowed  to  remain 
about  twenty  hours.  The  solution  is  next  drained 
slowly  off  into  its  sump,  allowing  four  hours.  A 
weak  solution  (9  pounds  of  0.05%)  is  next  run  on, 
a  portion  being  drawn  off  immediately  and  run  into 
the  strong  solution  sump  until  equal  in  quantity  to 
the  strong  solution  originally  applied.  Contact  is 
again  allowed  for  twenty  hours  with  four  hours  for 
drawing  off.  A  water  wash  of  10  pounds  is  finally 
run  on  and  in  an  hour  or  two  as  much  is  drawn  off 
into  the  weak  solution  sump  as  there  was  weak  solu- 
tion added.  The  water  wash  may  then  be  allowed 


54 


TESTING  FOR  PROCESSES 


Solution 


Leacfrzng  Tew  'A. 

Zcacfa'Hy  Tank 

©>  20" 

Zeacfa'xy  Tan  A. 

t,8                       A 

<si 

"V 

Sump. 


Fig.  10.     LABORATORY  CYANIDE  PLANT 


to  run  off  immediately,  but  slowly,  so  as  to  take 
perhaps  twenty-four  hours  from  the  time  of  apply- 
ing. The  ore  residue  is  next  assayed,  and  the  sump 


CYANIDATION  55 

solution  passed  through  the  zinc  boxes  or  some  other 
form  of  precipitation  apparatus.  If  zinc  shavings 
are  used,  the  metal  is  recovered  by  dissolving  the 
shavings  in  acid,  collecting  the  residue  on  a  filter, 
and  then  the  whole  metal  value  recovered  as  de- 
scribed under  'Bullion  Refining.'  The  solutions  are 
assayed  for  gold  and  silver,  before  and  after  passing 
through  the  zinc  boxes.  This  will  give  the  efficiency 
of  the  precipitation. 

DOUBLE  TREATMENT 

A  modification  of  cyanide  process  is  to  give  part 
treatment  in  one  vat,  and  then  to  transfer  the  mate- 
rial, after  draining,  to  a  second,  where  the  leaching 
is.  completed.  There  are  two  advantages  in  this 
system  of  working:  (1)  in  turning  over  the  whole 
mass  during  transfer  any  closely  packed  unleach- 
able  portions  in  the  first  vat  may  be  broken  up  and 
distributed  among  the  rest  of  the  sand;  (2)  the 
material  partly  treated  and  moistened  with  cyanide 
solution  is  exposed  to  contact  with  the  oxygen  of 
the  air  during  the  removal  from  one  vat  to  the 
other.  Usually  in  the  first  vat  an  alkaline  wash,  a 
weak  cyanide  solution  and  a  medium  solution  or  part 
of  the  strong  solution,  are  passed  through  the  mate- 
rial in  order  that  a  sensible  amount  of  cyanide  solu- 
tion may  be  present  at  the  time  of  transfer.  Then, 
in  the  second  vat,  the  remainder  of  the  strong  solu- 
tion is  added,  followed  in  the  usual  way  by  the  weak 
solution  and  water  wash.  There  is  no  doubt  what- 
ever as  to  the  improvement  in  percolating  qualities 
caused  by  the  transfer  from  one  vat  to  the  other,  in 


56  TESTING  FOR  PROCESSES 

the  case  of  ordinary  mixed  or  unclassified  material; 
and  even  with  the  sand  product  of  the  present  clas- 
sification methods  there  is  a  sensible  advantage  in 
this  respect,  for  it  is  found  that  these  sands  occupy 
a  larger  space,  being  sometimes  20%  more  in  the 
second  vat  than  in  the  first.  As  regards  oxidation, 
there  must  be  also  some  slight  advantage,  particu- 
larly when  dealing  with  unclassified  sand. 

PRECIPITATION  BY  MEANS  OF  ZINC 

When  a  positive  metal  having  a  difference  of  po- 
tential at  two  parts  of  its  surface,  is  immersed  in  a 
salt  of  a  less  positive  one,  the  dissolving  part  becomes 
the  anode  and  the  less  soluble  part  the  cathode ;  and 
precipitation  of  the  less  positive  metal  proceeds  as 
if  the  current  were  applied  from  an  external  source. 
If  the  salt  is  a  complex  one  as  KAu(CN)2,  the  action 
may  be  said  to  take  place  as  follows.  Let  it  be 
assumed  that  the  precipitating  metal  is  zinc,  which 
contains  lead  as  an  impurity,  and  the  solution  potas- 
sium cyanide  containing  some  KAu(CN)2.  Currents 
are  generated  by  the  dissolving  action  of  the  zinc 
and  flow  through  the  solution  from  the  zinc  to  the 
lead.  In  the  paths  of  these  currents,  molecules  of 
KAu(CN)2  are  carried  by  diffusion  and  their  ions 
are  directed,  the  Au(CN)2  going  to  the  zinc  and  the 
K  to  the  lead,  where  they  simultaneously  give  up 
their  charges  and  become  atoms  or  molecules.  The 
gold  deposits  on  the  part  of  the  zinc  where  the  cur- 
rent enters  and  the  (CN)2  is  left  free  to  combine 
with  KOH  or  water.  The  K  ions  give  up  their  charge 
in  the  same  way  as  the  lead  and  becomes  an  atom 


CYANIDATION  57 

forming  for  the  moment  an  anode,  while  the  lead  is 
the  cathode.  Gold  is  then  deposited  on  the  lead  as 
the  solution  diffuses  to  that  electrode.  Thus,  when 
a  zinc-lead  couple  is  employed,  gold  deposits  both 
at  the  zinc  and  the  lead  pole,  whereas  when  we  apply 
a  current  from  an  external  source,  gold  is  deposited 
permanently  only  at  the  cathode.  At  the  moment 
the  atom  of  gold  is  deposited  on  the  zinc  its  tendency 
is  to  redissolve,  but  a  new  positive  or  anode  part 
of  the  zinc  takes  this  atom  as  a  negative  electrode 
or  cathode  on  which  to  deposit  potassium  atoms. 
The  potassium,  as  it  separates,  keeps  up  local  cur- 
rents that  precipitate  further  quantities  of  gold  as 
the  solution  diffuses,  and  thus  the  deposit  thickens. 
The  proportion  of  gold  actually  precipitated  by  the 
primary  action  of  the  zinc  must  necessarily  be  small, 
as  the  surface  soon  becomes  covered  with  gold,  and 
the  chief  amount  must  be  obtained  by  the  secondary 
action  of  the  potassium.  In  the  case  of  precipitation 
from  very  weak  solutions  as  those  used  in  the  treat- 
ment of  slime,  the  use  of  the  zinc-lead  couple  ob- 
tained by  dipping  the  zinc  shavings  in  lead  acetate 
will  give  very  satisfactory  results.  With  strong 
solutions  the  precipitation  by  zinc  alone  is  efficient, 
and  the  use  of  the  couple  is  not  necessary.  With 
zinc  shavings  alone,  all  authorities  are  of  the  opinion 
that  strong  solutions  give  uniformly  good  results, 
while  weak  solutions  give  results  that  are  more  or 
less  erratic. 

Consumption  of  Zinc. — The  actual  amount  of  zinc 
dissolved  during  the  precipitation  is  a  matter  that  is 
not  easy  to  determine,  but  it  is  certainly  many  times 


58  TESTING  FOR  PROCESSES 

in  excess  of  the  theoretical  quantity  required.  If 
the  whole  of  the  zinc  dissolved  were  used  to  precip- 
itate the  gold  or  silver  in  the  solution,  1  ounce  should 
deposit  6  of  gold  or  3  of  silver,  and  this  is  about 
what  would  be  obtained  in  a  concentrated  solution 
of  KAu(CN)2  or  KAg(CN)2.  In  actual  practice  it 
requires  from  seven  to  ten  times  the  theoretical 
amount  of  zinc,  which  for  the  most  part  is  wasted 
in  decomposing  water  and  liberating  hydrogen. 

Zinc  Box  Operations. — The  gold-bearing  solution 
is  allowed  to  flow  through  the  box  at  as  uniform  a 
rate  as  practicable.  The  most  suitable  rate  of  flow, 
which  varies  with  the  sectional  area  of  the  column 
of  zinc,  with  the  area  of  the  zinc  surface,  and  with 
other  factors,  is  found  by  trial.  With  a  little  prac- 
tical experience  the  operator  gets  a  rough  general 
idea  as  to  whether  precipitation  is  effective  or  not. 
If  the  precipitate  is  loose,  of  a  black  or  dark  brown- 
ish color,  it  may  be  taken  for  granted  that  the  gold 
is  precipitating  satisfactorily.  If,  on  the  other  hand, 
the  deposit  is  firm  and  light  colored,  it  is  probable 
that  the  sump  solutions  will  assay  high.  It  is  usual 
then  to  take  the  precaution  of  adding  a  few  pounds 
of  cyanide  in  lumps  or  of  caustic  alkali  to  the  head 
of  the  zinc  box  and  thus  temporarily  increase  the 
action  of  the  zinc.  The  light  color  then  disappears 
and  the  deposit  should  become  loose.  The  light- 
colored  precipitate  is  not  always  a  sign  of  poor  pre- 
cipitation, as  it  may  be  caused  by  aluminous  or 
other  matter  precipitated  from  the  solution,  but  it 
is  generally  advisable  to  keep  any  such  precipitates 
at  a  minimum.  If  this  cannot  be  done  by  temporar- 


CYANIDATION  59 

ily  increasing  the  strength  of  the  solution,  it  is  ad- 
visable to  clean  up  often. 

Copper  in  Solution.— Much  difficulty  has  often  been 
experienced  by  the  presence  of  copper  in  solution. 
Zinc  in  a  cyanide  solution  precipitates  copper,  but 
the  potential  difference  between  the  two  metals  is 
small  compared  with  that  between  zinc  and  gold,  or 
zinc  and  silver.  The  effect  of  this  is  that  a  solution 
which  would  just  precipitate  a  loose  black  deposit 
of  gold  would  not  be  at  all  suitable  for  throwing  down 
a  similar  mass  of  copper.  If  copper  were  present 
in  such  a  solution  it  would  be  precipitated  as  a  firm 
metallic  coating,  such  as  would  raise  the  decomposi- 
tion point  of  the  solution  above  that  which  the  po- 
tential of  the  zinc  was  capable  of  effecting,  and  then 
the  action  would  fall  off.  Should  the  solution  be 
sufficiently  strong  in  salts  to  deposit  the  copper  in 
a  loose  form,  both  gold  and  copper  precipitate  to- 
gether ;  but  as  the  solution  pressure  of  the  copper  is 
higher  than  that  of  gold,  a  larger  porportion  of  the 
former  is  re-dissolved  and  the  net  result  is  a  poorer 
precipitation.  The  strength  of  solution  required  to 
produce  this  loose  deposit  of  copper  is  not  often  met 
in  present  treatment  of  gold  ores,  and  it  is  therefore 
advisable,  from  an  economic  point  of  view,  when  cop- 
per is  troublesome,  to  resort  to  other  methods  of  pre- 
cipitation, such  as  the  electrical  process.  When  cop- 
per is  known  to  be  present  in  the  ore,  care  should  be 
taken  to  make  exhaustive  quantitative  tests  before 
deciding  to  employ  zinc  as  a  precipitant. 

Electrical  Precipitation. — When  an  electric  cur- 
rent is  applied  to  a  cyanide  solution  containing  gold 


60  TESTING  FOR  PROCESSES 

or  silver,  the  K  ions  travel  toward  the  cathode  and 
the  Au(CN)2  or  Ag(CN)2  toward  the  anode.  Hence 
it  is  important  to  overbalance  this  influence  of  the 
current  by  giving  motion  to  the  solution  so  as  to 
bring  the  gold  and  silver-bearing  ions  in  contact 
with  the  cathode.  Some  of  the  solution,  however, 
must  get  to  the  anode  and  is  disassociated  as  atoms, 
and  this  accounts  for  the  gold  found  with  the  ferric 
oxide  attached  to  the  iron  plates  when  they  are  used 
as  anodes. 

It  has  been  found  that  to  obtain  a  good  adherent 
deposit  in  electroplating,  the  current  must  be 
roughly  proportional  to  the  quantity  of  gold  per 
unit  of  solution.  For  a  current  of  10  amperes  per 
square  foot,  the  poorest  solution  used  contains  50 
ounces  of  gold  per  ton,  and  this  is  equivalent  to  0.01 
ampere  per  square  foot  for  each  0.05  ounces  of  gold 
in  solution. 

Siemens  &  Halske  Process. — In  this  process  the 
anodes  are  of  sheet-iron  from  %  to  */£  inch  in  thick- 
ness, sewed  up  in  covers  of  hessian  or  other  textile 
material  to  prevent  short  circuiting.  The  plates  are 
connected  to  the  positive  pole  of  the  dynamo.  The 
cathodes  consist  of  lead  foil  0.0032  inch  thick,  weigh- 
ing about  0.19  pound  per  square  foot.  These  are 
connected  to  the  negative  pole  of  the  dynamo.  The 
average  current  used  is  0.04  amperes  per  square  foot 
of  anode  with  a  cathode  surface  of  50  to  150  square 
feet  per  ton  of  solution  per  24  hours,  according  to 
the  conditions. 

Butters  Process. — This  differs  from  the  Siemens  & 
Halske  process  in  that  insoluble  anodes  of  peroxi- 


CYANIDATION  61 

dized  lead  are  used.  They  do  away  entirely  with 
the  sludge  due  to  the  dissolution  of  iron  anodes. 
They  also  make  it  possible,  by  the  use  of  a  high  cur- 
rent density,  to  precipitate  the  bullion  as  a  pulver- 
ulent deposit  which  can  be  recovered  by  wiping  the 
cathode,  thus  obviating  the  necessity  of  destroying 
the  cathode  to  separate  the  metallic  deposit,  as  is 
necessary  in  the  Siemens  &  Halske.  The  ordinary  tin 
plate  of  commerce  is  used  as  a  cathode,  and  the  cur- 
rent density  of  0.28  to  0.56  amperes  per  square  foot 
of  anode  is  used. 

TREATMENT  OF  CONCENTRATE 

The  cyanide  process  is  now  applied  in  many  cases 
to  materials  which,  a  few  years  ago,  would  without 
question  have  been  treated  by  roasting  and  subse- 
quent chlorination.  This  is  true  especially  of  pro- 
ducts of  close  concentration,  which  are  often  cya- 
nided  raw,  but,  if  rich  enough,  are  preferably 
roasted.  Cyanidation  was  at  first  applied  to  these 
refractory  materials  only,  which  were  then  treated 
by  agitation,  alternating  with  percolation.  From  a 
purely  technical  point  of  view,  this  is  undoubtedly 
the  quickest  and  most  efficient  method  of  extracting 
the  gold  from  such  material,  whether  raw  or  roasted. 
But  the  expense  of  continuous  or  repeated  agitation 
of  coarse  particles  of  high  specific  gravity  limits  the 
application  of  this  method  to  rich  material.  The 
general  practice  is  to  crush  the  whole  of  the  material 
very  fine  and  to  treat  it  as  slime  by  agitation  and 
filter  pressing. 

In  treating  slime  by  cyanidation  there  are  two 


62  TESTING  FOR  PROCESSES 

main  difficulties:  (1)  that  due  to  the  viscosity  of 
the  pulp  which  diminishes  the  dissolving  power  of 
the  solvent  for  gold  and  silver;  (2)  due  to  the  pres- 
ence of  substances  capable  of  combining  with  ab- 
sorbed oxygen.  To  overcome  these  difficulties  the 
viscosity  of  the  pulp  is  diminished,  air  is  supplied 
from  an  external  source,  and  the  pulp  kept  agitated 
while  in  the  vats. 

SLIMING  TESTS  FOR  5  TO  20  POUNDS  OF  ORE 

Preparation. — The  sample  of  the  ore  is  ground  in 
the  coffee  mill  to  30  or  40  mesh  and  then  introduced 
into  the  pan  grinders  used  for  silver  amalgamation 
where  it  is  made  into  a  thick  pulp  with  water  and 
reduced  to  the  required  fineness.  Lime  is  added  as 
previously  determined.  A  dilute  cyanide  solution 
of  0.005  to  0.10%  strength  is  generally  used  instead 
of  water.  Use  approximately  one  pound  of  solution 
per  pound  of  ore.  The  slime  is  agitated  in  an  ap- 
paratus shown  in  Fig.  11.  When  agitating  slime 
alone  use  six  parts  of  solution  to* one  part  of  slime. 
When  both  sand  and  slime  are  present,  a  thicker 
pulp  must  be  used  according  to  percentage  of  slime 
present. 

Use  of  Vacuum  Leaf  Slime  Filter. — Run  the  slime 
into  the  filter-tank  when  the  agitation  is  finished 
and  start  the  air  pump  connected  with  the  filter- 
frame.  Note  the  vacuum  and  the  time  required  to 
build  up  a  cake  of  slime  on  the  filter,  one  inch  thick. 
Always  keep  the  filter-frame  covered  with  solution. 
When  the  required  slime  cake  has  been  built  up, 


CYANIDATION 


63 


64  TESTING  FOR  PROCESSES 

remove  to  the  weak  solution  and  draw  in  an  amount 
of  solution  equal  to  twice  the  moisture  in  the  cake. 
Measure  the  amount  of  solution  drawn  in  by  the 
lowering  of  its  level  in  the  tank.  Remove  to  the 
wash  water  and  draw  in  one-half  as  much  as  of  the 
weak  solution.  Remove  the  filter,  shut  off  the  vac- 
uum, turn  on  the  water  pressure  and  discharge  the 
adhering  slime  cake. 

Note  the  time  required  for  these  operations,  the 
area  of  the  filter-frame,  and  the  weight  of  dry  slime 
treated.  Assay  the  discharged  tailing.  From  the 
above  data  compute  the  filter  area  required  for  a 
given  tonnage  (one  ton  per  24  hours),  the  percent- 
age of  extraction,  consumption  of  cyanide,  and  other 
calculations  pertaining  to  the  commercial  success  of 
the  test. 

CYANIDATION  OF  SILVER  ORES 

The  success  of  the  application  of  cyanidation  to 
treatment  of  silver  ores  is  greatly  dependent  upon 
the  condition  in  which  the  silver  exists  in  the  ore ; 
that  is,  whether  it  is  soluble  or  insoluble  or  in  com- 
binations with  minerals  that  have  injurious  effects 
upon  cyaniding.  The  forms  of  silver  practically  in- 
soluble or  sparingly  soluble  in  cyanide  are,  coarse 
native  silver,  ruby  silver  (Ag3AsS3  or  AgSbS3),  and 
stephanite  ( AgSbSJ .  These  last  two  are  more  read- 
ily soluble  in  mercuric  cyanide.  The  silver  minerals 
readily  soluble  in  cyanide  are,  the  chloride,  bromide, 
and  chloro-bromide  of  silver.  The  tests  for  cyani- 
dation of  silver  ore  are  practically  the  same  as  with 
a  gold  ore  with  the  following  exceptions;  the  time 


CYANIDATION  65 

of  treatment  is  usually  greatly  lengthened,  stronger 
solutions  are  required  (0.25  to  0.75%  KCN),  sliming 
is  generally  necessary  with  a  thorough  and  prolonged 
agitation  and  oxygenation  of  the  pulp.  A  greater 
consumption  of  cyanide  must  be  expected  on  account 
of  the  larger  amount  of  metal  extracted  as  compared 
with  gold.  Concentration  may  often  precede  the 
cyanide  treatment  to  advantage.  Rebellious  ores 
might  be  given  a  chloridizing  roast,  as  a  plain  roast 
invariably  interferes  very  seriously  with  silver  ex- 
traction. The  increased  extraction  due  to  the  roast 
must  warrant  its  high  cost  and  loss  of  silver. 


CHLORIDIZING  ROAST  OF  SILVER  ORES 

Principles  and  Reactions. — Chloridizing  roasting 
is,  generally  speaking,  a  necessary  preliminary  to 
the  treatment  by  wet  processes  (amalgamation  or 
lixiviation)  of  all  the  silver  ores  in  which  that  metal 
occurs  as  a  sulphide,  sulphantimonide,  or  a  sulph- 
arsenide,  as  well  as  those  in  which  silver  sulphide 
occurs  not  merely  mingled  but  isomorphously  asso- 
ciated with  galena,  blende,  or  other  sulphides.  Ores 
for  chloridizing  require  to  be  crushed  from  16  to 
40  mesh  preparatory  to  roasting.  When  the  silver- 
bearing  mineral  is  brittle  and  decrepitates  easily 
with  heat,  the  ore  only  requires  crushing  to  16  mesh, 
but  dense  ores  which  do  not  have  this  property,  espe- 
cially those  containing  much  blende  or  galena,  re- 
quire to  be  crushed  to  40  mesh  to  secure  good  results. 
The  chloridizing  roast  aims  to  convert  as  much  as 
possible  of  the  silver  present  into  chloride,  and  of 
the  other  metals  into  their  oxides;  the  agency  em- 
ployed being  common  salt. 

The  ores  are  usually  divided  into  three  groups : 
(1)  those  containing  more  than  8%  sulphur,  which 
usually  require  a  preliminary  roast  before  adding 
the  salt;  (2)  ores  containing  3  to  S%  sulphur; 
(3)  ores  containing  less  than  3%  sulphur.  These 
usually  require  the  addition  of  sulphur,  as  iron  py- 
rite,  in  order  to  decompose  the  salt  and  liberate  the 
chlorine. 

The  percentage  of  salt  used  varies  from  10  to  15% 


CHLORIDIZING  SILVER  ORES  67 

and  can  be  determined  best  by  trial.  With  some  ores 
it  is  found  better  to  add  salt  at  the  beginning ;  with 
others  near  the  end  of  the  roast.  Ores  containing  no 
As,  Sb,  Pb,  or  Ca  are  generally  best  treated  by  add- 
ing salt  at  the  beginning.  If  arsenic  volatilizes  as  a 
sulphide  the  loss  of  silver  seems  to  be  less  than  when 
it  volatilizes  as  a  chloride.  Lead  and  lime  should, 
if  possible,  be  kept  as  sulphates,  for  otherwise  they 
are  great  consumers  of  chlorine.  If  an  ore  carries 
sulphides  of  lead,  zinc,  copper,  and  iron,  it  is  roasted 
at  a  just  visible  red,  in  order  to  form  sulphates. 
Sulphates  of  iron  and  copper  will  decompose  salt 
with  the  liberation  of  chlorine  and  the  other  two 
will  not. 

2NaCl  +  2FeS04  +  30  =  4C1  +  2Na2S04  +  Fe2O3. 

It  is  therefore  advisable  to  form  FeS04,  CuS04,  and 
as  much  sulphate  and  oxide  of  lead  and  zinc  as  pos- 
sible, and  yet  not  decompose  the  FeS04  and  the 
CuS04.  For  this  reason  the  roast  is  carried  on  slowly 
and  at  a  low  temperature  until  the  required  stage  is 
reached.  The  salt  may  then  be  added.  The  roast 
is  ordinarily  finished  at  a  cherry-red  heat.  The  chlo- 
rine liberated  together  with  HC1  formed  will  chlo- 
ridize  the  ore,  as  below : 

Steam  (H20)  +  CuCl2  =  CuO  +  2HC1 

2HC1  +  Ag2S  =  2AgCl  +  H2S 

8C1  +  Ag2S  +  4H20  =  2AgCl  +  H2S04  +  6HC1 

Careful  note  should  be  taken  of  the  duration  of 
the  roast,  heat  used,  and  at  what  time  the  salt  was 
added. 

Testing  an  Ore. — Crush  the  ore  so  as  to  pass 
through  a  30  or  40-mesh  screen.  Sample  carefully 


68  TESTING  FOR  PROCESSES 

and  determine  the  total  gold  and  silver  content. 
Weigh  out  from  300  to  500  grams  of  the  ore  and 
roast  in  an  iron  dish  in  the  gas  oven  furnace.  The 
time  of  adding  the  salt  and  the  amount  used  is 
determined  by  the  sulphur  content  of  the  ore  as 
previously  given.  Roast  slowly  and  carefully  (from 
30  minutes  to  two  hours),  with  a  frequent  stirring 
to  prevent  caking,  and  finish  at  a  cherry-red  heat. 
In  practice  the  chlorination  is  often  increased  by 
withdrawing  from  the  furnace  and  allowing  the  ore 
to  stand  on  a  cooling-floor  from  24  to  60  hours  and 
then  moistening  the  cooling  heaps  with  water,  by 
which  means  an  additional  chlorination  of  3  to  6% 
is  often  reached. 

After  roasting  weigh  carefully  to  determine  loss 
and  then  pass  through  a  30-mesh  screen.  Sample 
carefully,  take  enough  ore  from  the  following  tests, 
and  grind  it  in  an  agate  or  porcelain  mortar  so  as 
to  pass  through  a  100-mesh  screen.  If  iron  recep- 
tacles were  used  for  grinding  the  following  reaction 
might  occur :  2AgCl  +  Fe  =  FeCl2  +  2Ag.  The  re- 
sult would  be  material  not  readily  soluble  in  hypo- 
sulphite solution. 

Take  5  grams  and  assay.  Calculate  the  total 
weight  of  gold  and  silver  in  the  roasted  ore,  then 
the  difference  between  this  and  the  total  amount 
in  the  raw  ore  equals  the  gold  and  silver  loss  dur- 
ing the  roast.* 


*See  'Reducing  Factors'  for  reason  for  calculating  losses 
on  total  weights  of  silver  and  gold  instead  of  difference 
of  assay  (p.  88). 


CHLORIDIZING  SILVER  ORES  69 

Soluble  Salts. — (Excess  of  NaCl  used  and  all  chlo- 
rides and  sulphates  soluble  in  water,  including  AgCl 
soluble  in  the  NaCl  solution,  if  an  excess  has  been 
used  in  the  roast).  Weigh  out  5  grams  of  the  roasted 
ore  and  leach  with  hot  water  by  decanting  through 
a  filter  paper.  Wash  the  ore  into  the  filter  with 
hot  water.  Dry,  burn  off  the  filter  paper  and  weigh. 
Loss  in  weight  equals  soluble  salts.  As  the  percent- 
age of  NaCl  used  in  the  roast  increases,  the  soluble 
salts  generally  become  more  abundant. 

Silver  as  Sulphate  and  Silver  Salts  Soluble  in 
Water  Containing  Salt  (NaCl). — Assay  the  whole 
residue  after  leaching  with  water.  The  difference 
between  this  weight  and  the  weight  of  the  button 
from  the  5  grams  of  roasted  ore  equals  the  silver 
salts  soluble  in  water  or  brine  solution.  Any  unde- 
composed  salt  left  in  the  ore  will  precipitate  the 
AgS04  as  chloride. 

Silver  Salts  Soluble  in  Hyposulphite. — Place  15 
grams  of  the  roasted  ore  in  a  beaker,  add  300  c.c.  of 
hot  water,  decant  the  water  through  a  filter;  add 
250  c.c.  of  5%  hypo  solution  (Na2S203  +  5H20),  al- 
low to  stand  on  ore  for  one-half  hour  at  125 °F., 
stirring  frequently. 

2AgCl  +  2Na2S2O3  =  2NaCl  +  2NaAgS208 
Decant  several  times  with  hot  water  through  the 
same  filter  as  first  used  and  finally  wash  with  a  lit- 
tle fresh  hypo.  Dry  the  residue,  weigh  and  assay 
the  whole  of  it.  The  difference  between  one-third 
the  weight  of  the  button  obtained  and  the  assay 
of  the  roasted  ore  equals  the  silver  salts  soluble  in 


70  TESTING  FOR  PROCESSES 

water  and  hypo.  Then  subtract  the  silver  salts  sol- 
uble in  water  and  get  silver  salts  soluble  in  hypo 
alone.  If  the  tailing  contains  a  quantity  of  gold 
that  will  pay  to  extract,  a  weak  solution  of  cyanide 
may  be  used  effectively  in  the  same  apparatus  by 
giving  the  ore  a  thorough  water  wash  following  the 
hypo  lixiviation  and  preceding  an  application  of  cy- 
anide solution.  The  gold  is  recovered  from  the  cy- 
anide solution  in  the  usual  way. 

Report. — Make  out  a  tabulated  report  of  results 
obtained  together  with  cost  data.  Compare  with 
cost  and  net  returns  by  shipping  direct  to  smelter. 


PAN  AMALGAMATION  FOE  SILVER  ORES 

Principles. — Amalgamation  of  ores  in  pans  is  used 
for  the  recovery  of  silver  from  free  milling  ores, 
those  in  which  the  silver  is  in  the  native  state  or  in 
the  form  of  a  chloride  or  bromide;  a  small  percen- 
tage of  the  sulphide  (argentite)  is  allowable.  If  the 
ore  is  not  free  milling,  it  must  first  be  given  a  chlori- 
dizing  roast,  as  already  described,  to  change  the  sil- 
ver-bearing sulphides,  arsenides  and  antimonites  to 
chlorides.  The  chief  object  in  pan  amalgamation  is 
to  keep  the  particles  of  ore  for  a  sufficient  length  of 
time,  in  contact  with  the  finely  divided  mercury 
globules  with  or  without  the  presence  of  certain 
reagents  to  aid  the  amalgamation.  The  pan  is 
used  instead  of  plates  as  the  ore  can  be  kept  in 
contact  with  the  mercury  for  a  much  longer  time  and 
facilitates  the  use  of  heat  and  chemicals. 

Interference  of  Various  Substances — Any  sub- 
stance which  tends  to  make  the  mercury  dirty  or 
coat  the  fine  globules  will  both  interfere  with  the 
amalgamation  and  increase  the  loss  of  mercury. 
Grease  in  any  form,  as  in  exhaust  steam,  is  fatal  to 
good  "results,  while  talc,  kaolin,  and  other  hydrated 
silicates  of  magnesia  and  alumina  act  similarly  by 
coating  the  globules  of  mercury  and  preventing  con- 
tact between  them.  Lead  ores  interfere  by  'sicken- 
ing' the  mercury.  Under  favorable  circumstances, 
the  lead  ores,  when  in  the  chloride  form,  are  rapidly 
amalgamated,  forming  a  pasty,  rapidly  tarnished 
amalgam,  which  is  very  easily  floured  and  thus  causes 


72 


TESTING  FOR  PROCESSES 


Fig.  12.     PAN  AMALGAMATOR  AND  GRINDER 


PAN  AMALGAMATION  73 

a  large  loss  of  mercury.  The  lead  chloride  may  be 
changed  to  the  sulphate,  which  will  not  readily  amal- 
gamate, by  the  addition  of  copper  sulphate  to  the 
charge,  as  below: 

PbCl2  +  CuS04  =  PbS04  +  CuCl2 

Arsenical  and  antimonial  ores  are  always  difficult 
to  work  as  the  mercury  not  only  becomes  'floured' 
but  also  *  sick '  and  '  dirty, '  so  that  it  will  not  readily 
amalgamate.  In  this  case  the  whole  stock  of  mer- 
cury should  be  frequently  cleaned  with  nitric  acid 
and  retorted.  Manganese  oxides  froth  in  the  pan 
and  give  rise  to  a  high  loss  of  mercury  and  a  low 
percentage  of  extraction. 

Addition  of  Chemicals. — In  the  case  of  some  free- 
milling  chloride  ores,  no  additions  of  chemicals  need 
be  made,  but  the  use  of  salt  has  the  effect  of  short- 
ening the  process.  In  the  case  of  all  refractory  ores, 
the  yield  is  increased  by  the  addition  of  both  salt 
and  copper  sulphate.  When  the  zinc  blende  and  py- 
rite  are  present  and  the  gangue  is  chalcareous,  some 
authorities  claim  that  salt  and  bluestone  should  not 
be  added.  Lime  consumes  bluestone,  forming  cal- 
cium sulphate,  and  the  sulphides  seem  to  flour  the 
mercury  in  the  presence  of  salt  and  perhaps  form 
chlorides.  With  special  ores,  other  additions  of 
chemicals  are  often  made.  Thus  with  ores  contain- 
ing oxide  or  carbonate  of  copper,  sulphuric  acid  in 
the  proportion  of  one  to  two  pounds  per  ton  may  ad- 
vantageously replace  all  or  part  of  the  copper  sul- 
phate. Lime  in  the  proportion  of  one  to  one-and- 
one-half  pounds  per  ton  is  frequently  employed  with 


7*  TESTING  FOR  PROCESSES 

ores  containing  partly  oxidized  pyrite.  Iron  borings 
in  quantities  up  to  20  Ib.  per  ton  are  frequent  addi- 
tions to  such  ores  as  corrode  the  pan  rapidly. 

Chemical  Reactions.  —  Many  reactions  take  place 
that  may  be  due  to  the  constituents  of  the  ore  or  may 
arise  from  chemicals.  The  following  are  some  of 
the  probable  reactions: 

CuS04  +  2NaCl  =  CuCl2  +  Na2SO4 
CuCl2  +  Ag2S  =  2AgCl  +  CuS 
CuO  +  H2S04  =  CuS04  +  H20 


2AgCl  +  Fe  =  FeCl2  +  Ag 

Ag2S  +  2Hg  =  Ag2Hg  +  HgS 
Procedure  for  Experiment.  —  Clean  all  rust  and 
dirt  from  the  pan  and  mullers  as  it  is  essential  to 
good  results  that  the  pan  bottoms,  mullers,  and  dies 
should  be  free  from  graphite  or  iron  rust.  Pour 
water  into  the  pan  to  the  depth  of  one  inch  and  set 
the  mullers  in  motion.  Now  slowly  add  three  or 
four  pounds  of  the  ore  which  has  been  previously 
sampled,  assayed  and  weighed.  Add  a  little  water 
from  time  to  time  to  keep  the  mass  in  a  pasty  con- 
dition. The  charge  should  finally  be  about  the  con- 
sistence of  a  thick  paint.  Care  should  be  taken  not 
to  have  the  pulp  too  thin  or  the  mercury,  when 
added,  will  sink  to  the  bottom  and  not  come  thor- 
oughly in  contact  with  all  the  ore  particles.  Add  the 
chemicals  suitable  for  the  ore  selected.  The  follow- 
ing may  be  taken  as  an  example:  salt  3l/2%  ;  copper 
sulphate  0.1  to  0.15  per  cent. 

Allow  the  mullers  to  rest  on  the  dies  and  grind  for 


PAN  AMALGAMATION  7S 

one-half  hour  or  until  all  the  lumps  have  been  ground 
and  the  ore  is  finely  divided.  Now  raise  the  mullers 
one-half  inch  from  the  dies  and  add  5  to  10%  mer- 
cury (weighed),  spraying  it  through  a  chamois  skin 
on  top  of  the  pulp.  If  the  pulp  is  of  the  right  con- 
sistence, the  mercury  will  be  found  disseminated  in 
fine  globules  throughout  the  mass.  With  the  aid  of 
a  Bunsen  burner,  heat  the  pulp  to  a  temperature  of 
175  to  185°  F.  and  maintain  this  temperature  until 
the  end  of  the  run. 

A.t  the  end  of  two  hours,  thin  out  the  pulp  with 
water  and  continue  the  agitation  for  15  minutes  to 
settle  out  the  mercury.  Pour  off  all  the  overlying 
pulp  possible  and  separate  the  mercury  and  amal- 
gam from  the  remainder  of  the  tailing  by  panning  or 
by  using  a  hydraulic  classifier.  The  tailing  should 
be  sampled  and  assayed  for  gold  and  silver.  The 
silver  is  recovered  from  the  amalgam  as  described 
under  'Retorting.'  Weigh  the  silver  and  mercury 
recovered. 

Report. — (1)  Extraction  based  on  assay  of  heads 
and  tailing;  (2)  extraction  based  on  actual  recov- 
ery; (3)  loss  of  mercury  in  pounds,  per  ton  of  ore; 
(4)  itemized  cost  of  the  operation. 


POT  ROASTING 

General  Principles. — 'Lime  roasting'  is  a  term 
proposed  by  W.  R.  Ingalls  for  the  operation  of  forc- 
ing air  under  pressure  through  a  mixture  of  galena 
and  lime  at  the  kindling  temperature,  with  the  ob- 
ject of  oxidizing  the  lead  and  the  sulphur  and  of 
fritting  or  fusing  the  charge.  It  may  be  considered 
as  a  limitation  of  the  broader  term  of  pot  roasting 
which  does  not  specify  the  use  of  lime  or  restrict 
the  process  to  galena  alone. 

At  the  present  time  there  are  two  views  as  to 
what  reactions  take  place  during  the  process,  but 
the  evidence  and  available  data  are  too  incomplete 
to  enable  a  definite  conclusion  to  be  reached.  Some 
facts,  may,  however,  be  stated.  There  is  clearly  a 
reaction  to  a  certain  extent  between  lead  sulphide 
and  lead  sulphate  as  in  the  reverberatory  furnace, 
because  prills  of  metallic  lead  are  to  be  observed  in 
the  lime-roasted  charge.  There  is  a  formation  of 
sulphuric  acid  in  the  lime  roasting,  upon  the  oxidiz- 
ing effect  of  which  Salvelsberg  lays  considerable 
stress,  since  its  action  is  to  be  observed  on  the  iron 
work  in  which  it  condenses.  Calcium  sulphate,  which 
is  present  in  all  the  process,  being  specifically  added 
in  the  Carmichael-Bradford,  evidently  plays  an  im- 
portant chemical  part,  because  not  only  is  the  sul- 
phur trioxide  expelled  from  the  artificial  gypsum, 
but  it  is  also  to  a  certain  extent  expelled  from  the 
natural  gypsum  in  the  Carmichael-Bradford  process. 
Further  evidence  that  lime  does  indeed  play  a  chem- 


POT  ROASTING  77 

ical  part  in  the  reaction  is  presented  by  the  phenom- 
ena of  lime  roasting  in  clay  dishes  in  the  assay  muf- 
fle, wherein  the  air  is  not  blown  through  the  charge, 
which  is  simply  exposed  to  superficial  oxidation  in 
ordinary  roasting.  The  desulphurized  charge  dropped 
from  the  pot  is  much  below  the  temperature  of 
fusion,  while  even  in  the  interior  pyrite  and  zinc 
blende  are  completely  oxidized.  This  at  least,  in- 
dicates atmospheric  action. 

Experimental  Work. — Take  20  Ib.  of  galena  con- 
centrate and  crush  through  a  10-mesh  screen.  Sam- 
ple carefully  and  assay  for  FeO,  CaO,  MgO,  Si02 
and  S.  Make  up  a  charge  of  ore  20  Ib.,  with  lime- 
stone or  quartz  as  required  to  make  a  singulo-sili- 
cate  slag,  and  enough  water  to  make  the  whole 
charge  of  the  consistence  of  brasque.  The  roasting 
pot  is  now  warmed  by  burning  a  layer  of  charcoal 
on  the  grate,  with  a  gentle  air-blast.  A  layer  of 
coarse  limestone  is  placed  over  the  grating  as  a  safe- 
guard, then  followed  by  a  thin  layer  of  ignited  char- 
coal and  the  blast  turned  on  to  about  two  inches 
water  pressure.  As  soon  as  the  charcoal  is  burning 
freely  it  is  covered  with  a  second  layer  of  limestone, 
pea  size.  The  charge,  which  should  be  thoroughly 
mixed,  is  now  fed  gradually.  A  thin  layer  is  spread 
over  the  limestone  and  as  soon  as  it  becomes  ignited 
and  glowing  spots  begin  to  appear  on  the  surface, 
a  second  layer  is  added  and  so  on  until  the  pot  is 
filled.  The  hot  junction  of  a  thermo-electric  pyrome- 
ter is  buried  half  way  in  the  charge  and  readings  are 
taken  every  minute.  The  blast-pressure  is  now  raised 
to  about  six  inches  of  wrater  and  held  until  no  more 


78 


TESTING  FOR  PROCESSES 


fumes  pass  off  from  the  surface.    Any  new  holes  that 
form  should  be  immediately  filled  by  poking  down 


<?   C, 


Fig.  13.     POT  ROASTER 

the  charge  into  the  cavities.     When  the  charge  is 
cooled  down  somewhat,  it  is  dumped  out.    The  coarse 


POT  ROASTING  79 

limestone  is  picked  off  the  bottom,  the  slagged  and 
unslagged  portions  are  separated  and  weighed;  the 
slagged  portion  is  crushed  and  added  to  the  pulveru- 
lent part,  the  whole  sampled  down  and  assayed  for 
gold,  silver,  lead  and  sulphur. 

The  conditions  of  the  test  may  be  varied  to  show 
the  effect  of  adding  more  or  less  limestone  or  other 
flux,  high  or  low  blast,  on  the  loss  of  lead,  silver 
and  the  elimination  of  sulphur.  This  roast  may  also 
be  used  for  sulphide  ores  other  than  galena,  such  as 
pyrite  and  chalcopyrite. 

Report  the  losses  of  lead,  gold,  silver,  and  sulphur 
in  the  roast.  Loss  in  weight  of  the  charge.  Percen- 
tage of  slagged  and  unslagged  material.  Time  of 
roast,  temperatures  and  blast  pressures.  Cost  data. 


PARKES  PROCESS 

General  Principles. — The  Parkes  process  is  based 
on  two  facts  (1)  the  greater  affinity  of  silver  for  zinc 
than  for  lead;  and  (2)  the  insolubility  of  zinc-silver 
alloys  in  lead  which  is  already  saturated  with  zinc. 
The  process  of  desilverization  consists,  in  brief,  of 
stirring  1  to  2%  of  zinc  into  a  bath  of  molten  work- 
lead  or  base-bullion  heated  to  above  the  melting  point 
of  zinc  (415°C.),  and  allowing  it  to  cool,  when  a 
crust  or  scum  rises  to  the  surface,  containing  nearly 
all  the  silver.  A  repetition  of  the  process  with  a 
smaller  amount  of  zinc  leaves  a  lead  which  is  prac- 
tically free  from  silver,  or  not  containing  more  than 
0.2  to  0.3  oz.  per  ton. 

Zinc  has  an  even  greater  affinity  for  gold  and  cop- 
per, and  copper  alloys  so  readily  with  zinc,  it  is  im- 
portant to  purify  the  bullion  as  much  as  possible 
by  softening  and  dressing  before  the  desilverization 
is  commenced;  otherwise  a  poor  copper  crust  is 
formed  before  silver  can  be  extracted,  thus  wasting 
zinc  as  well  as  time.  It  is  noticeable  that  a  gold-cop- 
per crust  containing  all  but  the  minutest  trace  of 
gold  present,  is  formed  long  before  the  lead  begins 
to  take  up  the  zinc  to  any  considerable  extent,  the 
gold-zinc  and  copper-zinc  alloys  being  apparently 
almost  insoluble  in  zinc-free  lead,  whereas  it  is  im- 
possible to  produce  a  silver  crust  without  first  satu- 
rating the  lead  at  the  given  temperatures.  Bismuth 


PARKES  PROCESS  81 

does  not  interfere  with  desilverization,*  but  anti- 
mony in  the  proportion  of  0.1%  and  arsenic  in  even 
smaller  proportions,  not  only  retard  the  rising  of 
the  crusts  but  even  prevent  a  clean  separation  from 
the  underlying  lead.  On  a  commercial  scale  the  an- 
timony is  removed  by  dressing  at  a  bright  red  heat 
and  the  arsenic  is  oxidized  by  softening  8  or  9  hr. 
at  a  yellow  heat.  Nickel  and  cobalt,  rarely  present, 
enter  the  crust  like  copper  but  both  are  removed  in 
the  dross. 

All  crusts  are  separated  as  far  as  possible  from 
adhering  lead  by  liquation  before  being  further 
treated.  This  may  be  done  in  furnaces  or  in  pots. 
The  liquation  pot,  of  whatever  form,  must  be  heated 
very  gradually,  otherwise  some  of  the  zinc  crust  will 
be  dissolved  in  the  liquated  lead.  When  no  more 
lead  exudes,  the  crust  is  withdrawn  and  broken  up 
for  retorting  off  the  zinc.  No  matter  how  carefully 
this  be  done  the  liquated  lead  will  contain  30 
or  40%  more  silver  than  the  desilverized  lead.  In 
practice  it  is  put  back  into  the  next  charge  entering 
the  zinking  kettles. 

Procedure. — Take  about  20  Ib.  of  lead,  sample  and 
assay,  melt  down  in  a  crucible  very  slowly.  When 
the  lead  is  finally  melted,  skim  off  the  dross,  being 
careful  to  keep  the  lead  just  above  the  melting  point. 
This  dross  or  skimming  will  contain  nearly  all  the 
copper,  if  the  melt  has  been  slow  enough.  Raise  the 

*Blsmuth  is  an  impurity  detrimental  to  'corroding' 
market  lead.  This  impurity  is  not  eliminated  in  the  Parkes 
process,  but  it  may  be  removed  by  either  the  Pattison  or 
the  Betts  electrolytic  process. 


82  TESTING  FOR  PROCESSES 

temperature  am}  keep  the  material  at  a  good  red 
heat  until  a  dross  forms.  The  lead  is  to  be  tested 
from  time  to  time  until  the  antimony  is  removed  as 
shown  by  the  appearance  of  a  sample.  Before  the 
antimony  has  been  removed,  a  sample  of  the  bul- 
lion taken  in  a  ladle  will  work;  that  is,  small  par- 
ticles of  black  melted  skimmings  will  float  on  the 
surface  of  the  lead,  with  a  rotary  motion  which  re- 
sembles that  of  particles  of  grease  on  hot  water.  As 
the  softening  approaches  the  finishing  point  the 
globules  become  less  in  number  and  size,  a  coating  of 
yellow  litharge  forms  more  readily  on  the  molten  red 
hot  lead,  and  finally  no  more  globules  are  seen  and 
litharge  forms  quickly.  When  the  antimony  has  been 
removed,  skim  off  the  dross  formed  during  the  ope- 
ration. 

Raise  to  a  yellow  heat  and  cook  for  some  time  with 
a  free  access  of  air  to  oxidize  the  arsenic.  Pour  a 
sample  of  lead  into  a  mold  and  allow  to  cool  slowly, 
skimming  with  a  flat  wooden  stick.  If  there  is  no 
more  arsenic  and  antimony,  the  surface  of  the  lead, 
upon  solidifying,  will  assume  a  rich  indigo-blue  color, 
can  be  easily  scratched  with  the  fingernail,  and  will 
show  a  brilliant  lustre  on  a  fresh  incision.  When  ar- 
senic has  been  eliminated,  cool  down  to  a  red  heat 
and  again  dross.  Weigh  each  dross  as  removed. 

Cool  down  to  a  little  above  the  melting  point  of 
zinc  (415  °C.)  so  the  stick  of  wood  will  just  be  ignited 
by  the  lead.  Take  sample  No.  1  for  assaying  by  dip- 
ping out  with  a  scorifying  dish  held  in  a  pair  of  tongs. 
It  is  very  important,  when  adding  zinc,  to  have  the 
temperature  of  the  lead  only  above  the  melting  point  of 


PARKES  PROCESS  83 

zinc,  otherwise  the  zinc  will  oxidize  and  prevent  the 
operation  being  successful.* 

For  the  first  zinking  add  6  oz.  of  zinc  and  stir  in 
well  with  a  stick  of  wood,  cool  down  until  the  lead  is 
just  molten  and  thoroughly  dross  off  the  zinc  crust 
that  rises  to  the  surface.  This  dross  will  contain 
most  of  the  silver  and  all  of  the  gold.  Heat  up  again 
to  a  little  above  415  °C.  and  take  sample  No.  2.  In 
the  second  zinking  add  4  oz.  of  zinc  and  proceed  as 
in  the  first  zinking.  Reheat,  stir,  and  take  sample 
No.  3.  For  the  third  zinking  add  2  oz.  and  proceed  as 
before.  Take  sample  No.  4.  After  the  last  zinking, 
the  lead  is  again  raised  to  a  good  yellow  heat  until 
any  remaining  zinc  is  driven  off.  After  this  is  ac- 
complished, skim  and  pour  into  a  heated  mold,  al- 
low to  cool  and  weigh. 

In  treating  the  zinc  crusts  take  all  of  them,  place 
in  a  crucible  (See  Fig.  14),  heat  very  slowly  to  the 
melting  point  of  lead,  and  allow  the  excess  lead  to 
liquate  off  slowly.  Gradually  increase  the  tempera- 
ture so  that  all  the  free-lead  will  drain  off  (in  the 
meanwhile  agitating  the  crusts  with  a  rod  so  as  to 
break  up  large  pieces  and  help  work  off  the  lead), 
but  do  not  heat  so  high  that  the  zinc  will  distill  or  oxidize. 
Weigh  and  assay  the  liquated  lead.  The  dry  zinc 
crusts  are  placed  in  a  retort  and  heated  in  a  gas  forge 
and  the  zinc  fumes  caught  in  a  condenser.  When  all 
the  zinc  is  distilled,  the  remaining  'rich-lead'  is 
poured  into  a  bullion  mold,  and  the  silver  and  gold 

*Before  zinking  it  is  well  to  add  a  small  strip  of  aluminum 
and  stir  into  the  lead.  This  will  prevent  oxidation  during 
the  operation. 


84  TESTING  FOR  PROCESSES 

recovered  by  cupelling  off  the  lead  in  a  large  bone 
ash  cupel.  If  the  lead  contains  too  much  zinc  it  will 
not  readily  cupel  and  the  samples  should  first  be 
scorifed. 

Report. — Weigh  up  the  silver  and  gold  recovered, 
assay  samples  and  note  the  amount  of  silver  and  gold 


Fig.  14.     LIQUATOR  FOR  ZINC  CRUSTS 

left  after  each  dressing.  Account  for  all  the  gold 
and  silver  which  in  20  Ib.  of  base  bullion  should  be 
equal  to:  (1)  amount  of  Au  +  Ag  recovered;  plus 
(2)  amount  left  in  the  lead ;  plus  (3)  amount  of  losses 
due  to  cupellation,  etc.  The  latter  may  be  deter- 
mined by  difference  between  Au  +  Ag  button  actu- 


PARKES  PROCESS  85 

ally  recovered  and  the  calculated  recovery.  Calcu- 
late the  percentage  of  recovery,  cost  and  the  profit 
or  loss  based  on  the  recovery.  Give  weight  of  origi- 
nal lead  ingot,  weight  of  the  liquated  lead,  amount 
of  zinc  added  and  weight  of  dross  taken  from  the 
lead,  with  per  cent  of  dross  as  compared  with  the 
weight  of  the  base  bullion  taken. 


ZIERVOGEL  PROCESS  OR  THE  SULPHATIZING 
ROASTING  OF  COPPER  MATTE 

THE  GENERAL  PROCESS 

Principles. — This  very  simple,  though  delicate, 
process  depends  upon  the  fact  that  the  silver  in  cop- 
per mattes  is  converted  into  silver  sulphate  by  means 
of  the  sulphuric  anhydride  evolved  from  the  decom- 
position of  iron  sulphate  and  copper  sulphate,  which 
are  formed  during  the  roasting  from  the  iron  and 
copper  sulphides  present  in  the  mattes.  The  pro- 
cess consists  of  roasting  the  sulphides  with  an  ex- 
cess of  air,  so  that  they  are  oxidized  partly  to  sul- 
phates, which  are  then  heated  sufficiently  to  decom- 
pose the  sulphates  of  iron  and  most  of  the  sulphates 
of  copper,  while  the  silver  remains  sulphatized  as 
AgS04  and  may  be  dissolved  out  with  boiling  water 
to  be  subsequently  precipitated  on  metallic  copper. 
Simple  as  it  appears  this  process  is  extremely  diffi- 
cult to  execute,  for  it  requires  a  high  degree  of  skill 
to  seize  the  exact  period  when  the  iron  and  copper 
sulphates  are  converted  into  their  insoluble  higher 
oxides  and  when  none  of  the  silver  sulphate  is  de- 
composed. 

Procedure. — Grind  about  1%  lb.  of  matte  so  as  to 
pass  through  60  mesh.  Take  sample  No.  1  and  assay 
for  gold,  silver,  iron,  copper  and  sulphur.  Carefully 
note  the  amount  taken  for  the  roast.  Put  the  matte 
in  a  cast-iron  roasting  dish  and  place  on  the  hearth 
of  the  gas-oven  furnace  which  has  previously  been 


ZIERVOGEL  PROCESS  87 

lighted.  Heat  up  to  325  °C.,  as  indicated  by  a  ther- 
mo-electric pyrometer.  The  hot  junction  should  be 
placed  at  the  surface  of  the  ore.  At  this  stage  the 
sulphur  of  the  matte  begins  to  burn  and  the  tempera- 
ture will  then  rise  rapidly  owing  to  the  heat  of  com- 
bustion derived  from  the  burning  of  the  sulphur  in 
the  matte.  (S  +  20  =  S02  +  71000  calories.)  Stir 
continuously  to  prevent  caking.  Roast  for  about 
three-quarters  of  an  hour  between  450  and  550°C., 
reaching  the  latter  temperature  at  the  end  of  the 
period.  By  this  time  the  major  portion  of  the  sul- 
phur will  have  been  oxidized.  Take  sample  No.  2 
with  a  sampling  spoon. 

Now  the  heat  derived  from  the  combustion  of  the 
sulphur  will  decline  and  the  gas  and  blast  must  be 
increased.  Roast  for  1%  hr.,  gradually  increasing 
the  temperature  to  690  °C.  at  the  end  of  this  time. 
During  this  period  the  sulphates  will  be  formed. 
Take  sample  No.  3. 

The  composition  of  matte  is  Cu2S,  FeS,  Ag2S. 

At  570°  C.  ; 

FeS  +  30  =  FeO  +  SO2. 


S03  =  FeS04. 
Cu2S  +  40  =  2CuO  +  SO. 
Cub  -f  S03  =  CuS04. 

The  formation  of  the  copper  sulphate  is  a  gradual 
one  and  is  not  dependent  upon  the  action  of  the 
gaseous  products  of  the  dissociation  of  the  sulphate 
of  iron. 

^FeS04  +  heat  =  FeO  +  S03. 
CuO  +  S03  =  CuS04. 


88  TESTING  FOR  PROCESSES 

This  is  shown  by  the  roasting  of  an  iron-free  cop- 
per matte  in  which  the  rapid  accumulation  of  the 
copper  sulphate  at  temperatures  corresponding  to 
those  at  which  this  compound  appears  in  ordinary 
matte,  proves  that  iron  sulphate  was  not  essential 
to  the  sulphatizing  of  the  copper  compounds. 

Now  raise  the  temperature  gradually  till  at  the 
end  of  35  to  40  minutes  the  temperature  reaches 
not  higher  than  845°  C.,  at  which  temperature  the 
maximum  silver  sulphate  will  be  formed.  Above 
850°  C.  it  will  be  rapidly  broken  up.  Take  sample 
No.  4. 

At  the  end  of  this  roast  the  sample  should  show 
little  copper  sulphate  and  no  cuprous  oxide  because 
the  latter  will  precipitate  the  silver  from  the  silver 
sulphate  in  the  form  of  bright  metallic  scales.  This 
lower  oxide,  if  present,  must  be  re-oxidized  by  con- 
tinuing the  roasting  with  an  excess  of  air  before 
the  product  is  ready  for  leaching.  Test  by  throwing 
a  sample  into  a  beaker  of  hot  water  and  when  no 
bright  spangles  of  silver  appear  the  roast  is  com- 
plete. 

Now  raise  the  temperature  to  900°  C.  in  the  next 
15  minutes.  Take  sample  No.  5.  This  sample  will 
show  the  loss  of  silver  sulphate  due  to  overheating. 

Assay  the  samples  No.  2  to  5  for  copper,  iron, 
and  sulphates,  also  total  silver  content. 

Reducing  Factors.— The  above  results  are  not  in 
a  condition  for  direct  comparison.  During  the  roast- 
ing operation  much  oxygen  from  the  air  was  taken 
up  in  the  formation  of  the  sulphates  without  a  cor- 
responding loss  of  sulphur.  The  weight  of  the 


ZIERVOGBL  PROCESS  89 

charge  increases  as  the  sulphates  continue  to  rise 
in  amount,  and  as  these  compounds  are  decomposed. 
The  volume  also  increases  and  diminishes,  as  may 
be  readily  seen  in  the  roasting  dish  in  the  furnace. 
In  analyzing  the  samples  taken  at  successive  stages 
of  the  process,  the  same  weight  of  each  sample  was 
taken.  Because  of  the  rapidly  changing  weight  of 
the  roasting  charge,  the  amount  of  copper,  iron  or 
silver  sulphates  in  these  equal  weights  does  not  rep- 
resent the  relative  amounts  of  these  compounds  in 
the  charge  in  these  stages  of  the  process.  To  permit 
comparisons  to  be  made  it  is  necessary  to  ascertain 
quantitatively  the  weight  of  the  charge. 

The  total  amount  of  the  copper  will  not  change 
during  the  operation.  The  amount  determined  in 
equal  weights  of  the  respective  samples  will  vary 
through  considerable  limits.  From  the  varying  per- 
centages of  copper  present  in  the  different  samples 
the  change  in  weight  of  the  charge  is  ascertained. 
For  a  decrease  in  percentage  of  copper  an  increase 
in  weight  of  the  charge  must  have  taken  place,  and 
for  an  increase  in  the  copper  content  of  a  sample, 
a  diminution  in  the  weight  of  the  charge.  In  other 
words,  the  relative  weights  of  the  charge  at  different 
stages  are  in  inverse  ratio  to  the  copper  content  of 
equal  weights  of  sample  taken  at  these  stages.  Put- 
ting the  weight  of  the  original  matte,  before  roast- 
ing commenced,  as  unity,  the  weight  at  any  step  in 
the  operation  may  be  obtained  as  follows: 

I :  X : :  %  Cu  in  sample :  %  Cu  in  original  matte. 

Here  X  =  the  ratio  of  the  weight  of  the  charge  at 
the  time  the  sample  was  taken  to  the  weight  before 


90  TESTING  FOR  PROCESSES 

the  roasting  commenced.  By  multiplying  the  per- 
centages of  the  different  sulphates  in  each  sample, 
as  obtained  by  analysis,  by  the  reducing  factor  for 
that  sample  the  percentages  which  represent  relative 
amounts  of  these  sulphates  in  the  entire  charge  at 
the  time  the  sample  was  taken  may  be  obtained. 
A  table  of  corrected  results  may  be  thus  constructed. 

METHOD  OF  ANALYSIS 

Total  Copper. — Take  a  one-gram  sample  and  treat 
with  7  c.c.  nitric  acid  and  5  c.c.  of  sulphuric  acid  in 
a  flat-bottomed  flask,  until  white  fumes  come  off. 
Cool,  dilute  with  50  c.c.  of  distilled  water,  transfer 
to  a  graduated  flask,  add  15  c.c.  of  ammonia,  dilute 
to  a  known  volume,  by  preference  100  c.c.,  allow 
the  precipitate  of  hydroxide  of  iron  to  settle.  By 
means  of  a  pipette  take  off  an  aliquot  portion  of 
the  supernatent  blue  solution,  free  from  iron  pre- 
cipitate, and  titrate  for  copper  with  a  standard  solu- 
tion of  potassium  cyanide. 

Soluble  Sulphates  of  Iron  and  Copper. — Weigh 
out  a  2-gram  sample,  place  in  a  No.  3  beaker  with 
150  c.c.  of  distilled  water.  Heat  on  a  hot  plate  and 
boil  for  a  few  minutes.  Filter  into  a  No.  4  beaker, 
wash  the  residue  by  decanting  several  times  with 
hot  water  through  the  filter.  Add  a  strip  of  alumi- 
num two  inches  square  and  one-sixteenth  inch  thick, 
and. a  few  drops  of  sulphuric  acid.  Boil  on  a  hot 
plate  till  all  the  copper  is  precipitated  and  the  iron 
is  reduced,  using  a  drop  of  K4FeCy  on  a  spot  plate 
as  an  indicator.  Decant  and  wash  through  a  filter. 
Titrate  the  filtrate  with  a  standard  solution  of  potas- 


ZIERVOGEL  PROCESS  91 

slum  permanganate  for  the  iron  content  and  cal- 
culate as  FeS04.  Dissolve  the  copper  remaining  on 
the  filter  and  in  the  beaker  with  a  little  dilute  nitric 
acid,  wash  the  filter  with  water  into  the  beaker,  add 
ammonia  in  excess,  and  titrate  with  standard  potas- 
sium cyanide  solution;  calculate  as  CuS04. 

Silver  Present  as  Sulphate. — Take  0.2  A.  T.  and 
treat  with  hot  water  as  above.  If  the  copper  in  the 
solution  amounts  to  more  than  1  to  2%  of  the  matte, 
use  the  following  method.  To  the  solution  of  the 
silver  sulphate  add  NaCl  solution  of  about  normal 
strength  to  precipitate  the  silver  as  chloride,  care- 
fully avoiding  an  excess  of  the  NaCl.  Lead  acetate 
and  a  few  drops  of  sulphuric  acid  are  added.  Let 
stand  for  six  hours,  filter,  scorify  the  precipitate, 
and  cupel  for  silver.  Calculate  as  AgS04.  If  the 
copper  present  is  less  than  \%  the  silver  in  the 
hot  water  solution  may  be  determined  by  Volhards 
method.  For  total  silver  the  scorification  assay  may 
be  used. 

EE-SULPHATIZING  THE  METALLIC  SILVER 

Metallic  silver  resulting  from  the  decomposition 
of  the  silver  sulphate  by  excessive  heating,  exists 
in  a  finely  divided  condition  in  the  over-roasted 
matte.  It  may  be  re-sulphatized  by  the  sulphuric  an- 
hydride evolved  from  decomposing  ferrous  sulphate. 
No  doubt  other  decomposable  sulphates  would  ac- 
complish similar  results,  providing  their  dissociation 
occurred  at  a  sufficiently  high  temperature.  The  sul- 
phates of  sodium  and  potassium,  as  such,  do  not 
decompose  readily  by  heat  alone.  Ferrous  sulphate 


92  TESTING  FOR  PROCESSES 

is  a  convenient  compound  to  employ,  not  only  be- 
cause of  its  comparative  cheapness,  but  because  it 
begins  to  decompose  at  a  temperature  far  below 
that  to  which  silver  sulphate  can  be  safely  heated. 
Though  it  begins  to  decompose  below  600°  C.,  the 
decomposition  is  not  rapid  till  650  to  750°  C.  is 
reached.  If  a  mixture  of  over-roasted  matte  and 
ferrous  sulphate  be  rapidly  heated  to  a  temperature 
near  700°  C. — that  at  which  the  basic  sulphate  dis- 
sociates— there  will  be  little  cupric  sulphate  formed, 
but  the  sulphuric  anhydride  fumes  will  act  directly 
on  the  metallic  silver.  There  will  therefore  be  no 
necessity  of  raising  the  temperature  or  prolonging 
the  time  of  the  roast  to  decompose  the  sulphates  of 
copper,  thereby  again  running  the  risk  of  decom- 
posing the  silver  sulphate. 

Mix  a  portion  of  the  over-roasted  material  with 
5%  of  its  weight  of  dehydrated  ferrous  sulphate 
and  heat  in  the  roasting  dish  in  the  furnace  for 
20  to  30  minutes.  Allow  the  temperature  to  rise 
slowly  above  590°  C.  and  finally  raise  to  750°  C. 
Take  a  sample  and  determine  the  silver  present  as 
sulphate  and  note  the  percentage  of  silver  re-sulphat- 
ized  by  the  roast. 

In  practice  the  silver  is  recovered  by  leaching  the 
matte  with  hot  water  and  precipitating  the  silver 
from  this  solution  by  passing  over  copper  strips. 
The  copper  goes  into  solution  as  sulphate  which  is 
recovered  by  crystallization  or  by  passing  over 
scrap  iron. 

Report. — Plot  all  the  data  obtained  in  a  series  of 
curves,  with  time  for  ordinates,  and  temperatures, 


ZIERVOGEL  PROCESS  93 

percentages  of  silver,  copper  and  iron  as  sulphates, 
as  insoluble  compounds,  and  loss  of  silver  as  or- 
dinates. 


LIXIVIATION  OF  COPPER  OPvES 

Principles. — The  extraction  of  the  copper  in  the 
ore  by  lixiviation  is  effected  by  the  use  of  copper 
solvents  such  as  sulphuric,  hydrochloric,  and  sul- 
phurous (H2OS02)  acids,  and  by  air  oxidation  (Rio 
Tinto  method).  The  ores  suitable  for  direct  treat- 
ment are  the  oxidized  varieties  that  do  not  contain 
prohibitive  amounts  of  the  acid-consuming  carbon- 
ates of  lime,  magnesia,  alumina,  and  iron  oxides  or 
carbonates.  Sulphide  ores  after  being  roasted  with 
or  without  salt,  will  yield  a  satisfactory  extraction 
with  plain  water  leaching  followed  by  a  weak  acid 
wash.  The  method  might  be  successfully  combined 
with  the  ammonia-cyanide  process  by  removing  the 
excess  of  copper  over  one  per  cent  before  the  appli- 
cation of  the  ammonia  cyanide. 

Tests. — Take  a  sample  of  the  ore  and  assay  for 
gold,  silver,  and  copper.  Make  a  mineralogical 
examination  for  acid-consuming  carbonates  and  to 
determine  in  what  form  the  copper  exists.  If  the 
copper  is  present  as  a  sulphide,  the  ore  must  first 
be  subjected  to  a  roast  as  follows: 

Take  500  grams  of  the  ore,  place  in  a  cast-iron 
dish  and  roast  in  a  muffle  or  oven  furnace  where 
the  heat  may  be  readily  controlled.  The  object  of 
the  roast  is  to  decompose  the  sulphides  and  form 
the  maximum  amounts  of  soluble  copper  salts.  See 
*  Sulphatizing  Roast,'  page  86.  A  chloridizing 
roast  may  prove  more  effectual  than  a  plain  roast, 
as  the  precious  metals  are  "changed  into  more  soluble 


LIXIVIATION  95 

forms,  and  as  practised  in  the  Henderson  process* 
has  proved  very  successful.  If  a  chloridizing  roast 
is  used,  proceed  as  described  under  the  '  Chloridizing 
Koast  of  Silver  Ores',  page  66.  When  the  roast  is 
finished,  cool,  weigh,  and  assay. 

Take  50  grams  of  ore,  place  in  a  300-c.c.  beaker 
with  150  c.c.  of  water  and  agitate  at  intervals  for 
one-half  hour.  Decant  through  a  filter  and  wash 
with  150  c.c.  of  water.  Dry,  and  assay  for  gold, 
silver,  and  copper.  The  result  will  be  the  amount 
soluble  in  water  alone. 

Take  50  grams  of  ore,  treat  with  water  as  above, 
return  to  the  beaker,  add  150  c.c.  of  3  to  5%  sul- 
phuric or  hydrochloric  acid,  decant  through  a  filter, 
wash  with  150  c.c.  of  water,  dry  and  assay.  A  hot 
acid  solution  will  give  a  higher  and  more  rapid 
extraction  but  may  result  in  an  excessive  consump- 
tion of  the  acid.  The  metal  extracted  by  water 
washes  subtracted  from  the  above  results  will  give 
amounts  further  soluble  in  acid  due  to  the  presence 
of  basic  compounds  and  oxides. 

Consumption  of  Acid. — Take  20  grams  of  ore, 
wash  thoroughly  with  water  and  introduce  into  a 
wide  mouthed  bottle  with  40  c.c.  of  standard  acid 
solution;  agitate  for  30  minutes  and  then  filter  off 
20  c.c.  of  the  acid.  Titrate  with  a  standard  alkali. 
Calculate  the  consumption  of  100%  acid  per  ton  of 
ore. 

Report  the  extraction  of  gold,  silver,  and  copper. 

*See  Austin's  'Metallurgy  of  the  Common  Metals,'  2nd 
FJd.,  page  356. 


96  TESTING  FOR  PROCESSES 

Compare   with   net   returns   by   shipping   direct   to 
smelter. 


ELECTROLYTIC  PROCESSES 

GENERAL  STATEMENT 

The  quantities  of  metal  transported  by  a  certain 
current  in  a  certain  time  are  proportional  to  the 
atomic  weights  of  the  metal  divided  by  the  valency 
in  which  it  exists  in  the  solution.  This  is  known 
as  'Faraday's  law.'  Solutions  containing  the  salts 
of  the  metal  and  generally  with  free  acid  present 
are  almost  entirely  used  in  commercial  electrolytic 
refining.  Only  those  metals  that  do  not  dissolve 
from  the  anode  with  the  evolution  of  hydrogen  have 
been  successfully  refined  up  to  the  present.  When 
a  metal,  such  as  sodium  or  aluminum,  cannot  be 
successfully  treated  in  a  liquid,  a  fused  electrolyte 
is  used.  The  deposition  of  pure  metal  in  electrolytic 
refining  depends  upon  the  fact  that  each  metal  has 
its  own  electro-motive  force  of  solution.*  Metals 
lower  in  the  scale  than  the  principal  metal  present 
are  eliminated  as  metal  particles  in  the  anode  slime, 
and  metals  higher  in  the  series  are  eliminated  as 
salts  dissolved  in  the  solution  or  precipitated  from  it. 

The  transportation  of  a  pure  metal  from  one  pure 
anode  to  another  in  the  same  physical  condition,  re- 
quires little  power.  In  commercial  work  time  is  an 
important  factor,  as  the  capacity  of  the  plant  varies 
inversely  with  the  speed  of  working,  hence  a  bal- 


*The  difference  in  voltage  existing  between  a  metal,  and 
a  solution  containing  the  same  metal,  in  which  it  is  im- 
mersed. 


98 


TESTING  FOR  PROCESSES 


ance  must  be  found  between  an  increased  cost  for 
a  greater  output  against  lower  costs  for  power  for 
a  decreased  output,  per  vat,  in  the  same  time.  Since 


II 

1 

i* 

ELECTROLYTIC  PROCESSES  99 

power  is  such  an  item  of  expense  it  is  one  of  the 
first  considerations  to  find  an  electrolyte  of  as  high 
electric  conductivity  as  possible. 

The  following  electromotive  forces  of  solution  are 
practically  correct  for  fluosilicate  solutions,  but 
may  vary  a  few  hundredths  of  a  volt  for  different 
strengths  of  electrolyte  and  also  for  different  solu- 
tions : 

Volts. 

Zinc   +0.52 

Cadmium +  0.16 

Iron   +0.09 

Lead    —  0.01 

Tin —0.01 

Arsenic    —  0.40 

Antimony    —  0.44 

Bismuth     —  0.48 

Copper    —  0.52 

Silver   —  0.97 

Mercury    —  0.98 

ELECTROLYTIC  REFINING  OF  COPPER 

Principles. — The  process  depends  upon  the  fact 
that  when  copper-bearing  alloy,  cast  into  slabs,  is 
used  as  the  anode  of  an  electric  current,  an  acid 
solution  of  copper  sulphate  as  the  electrolyte,  re- 
fined copper  as  the  cathode,  and  a  current  of  suitable 
density  is  employed,  the  copper  will  be  dissolved  from 
the  anode  and  deposited  at  the  cathode,  while  the 
gold  and  silver  will  remain  at  the  anode  and  fall 
to  the  bottom  of  the  tank  in  the  form  of  mud.  At 
the  anode  S04  is  separated  and  a  corresponding 
amount  of  copper  is  dissolved ;  through  this  solution 
of  the  copper  the  electric  current  is  reinforced  by 


100  TESTING  FOR  PROCESSES 

an  amount  of  energy  that,  to  a  great  extent,  neu- 
tralizes the  back  electromotive  force.  In  the  process 
impurities  have  various  effects.  Gold,  silver,  and 
platinum  are  precipitated  as  metallic  particles  in  the 
slime.  Lead  becomes  a  sulphate.  Arsenic,  bismuth, 
and  antimony  go  into  solution  and  are  precipitated 
as  basic  sulphates,  but  as  the  electrolyte  becomes 
neutral  or  foul  the  arsenic  and  bismuth  are  partly 
carried  over  and  precipitated  at  the  cathode.  Tin 
dissolves  and  is  precipitated  as  a  basic  sulphate. 
However,  it  is  not  harmful,  but  improves  the  cath- 
ode deposit  and  diminishes  the  tension  of  the  bath. 
Iron,  zinc,  nickel,  and  cobalt  dissolve  and  impoverish 
the  bath  as  regards  both  copper  and  free  acid.  The 
neutralization  of  the  solution  has  many  prejudicial 
effects  upon  the  electrolytic  process.  The  solution 
becomes  a  poorer  conductor  and  foreign  metals  are 
deposited  at  the  cathode  along  Avith  cuprous  oxide. 
In  order  to  produce  good  copper  the  solution  must 
circulate  as  actively  as  possible,  because  copper  de- 
posited at  the  cathode  will  be  purer  and  more  mal- 
leable. If  this  solution  is  not  in  circulation  it  be- 
comes richer  in  copper  at  the  anode  and  poorer  at 
the  cathode;  the  poorer  solution  ascends  in  the 
bath,  thus  introducing  a  higher  resistance  in  the  cir- 
cuit. This  renders  it  possible  for  other  metals  or 
hydrogen  to  be  separated  along  with  the  copper. 
As  regards  the  relation  between  the  strength  and 
tension  of  current,  the  same  amount  of  copper  can 
be  obtained  with  a  given  power,  whether  the  voltage 
is  high  or  low,  as  long  as  the  current  density  is  the 
same;  i.  e.,  the  number  of  amperes  per  square  foot 


ELECTROLYTIC  PROCESSES  101 

of  anode  surface.  One  ampere  will  deposit  approx- 
imately one  ounce  of  copper  per  24  hours. 

The  electrolyte  is  purified  by  aerating;  boiling 
with  metastannic  acid ;  filtration  through  cuprous 
oxide ;  by  replacing  with  new  electrolyte  and  work- 
ing the  old  for  copper  and  sulphate  of  copper ;  or 
depositing  the  arsenic  by  taking  a  portion  of  the 
electrolyte  continuously,  sending  through  a  tank 
with  insoluble  lead  anodes  and  copper  cathodes, 
where  the  arsenic  and  excess  copper  are  deposited 
together  by  a  high  density  current.  Experience 
shows  that  the  liquid  most  suitable  for  electrolysis 
contains  per  gallon  0.5  pound  of  concentrated  sul- 
phuric acid  (H2SOJ  and  5  pounds  of  bluestone  con- 
taining 0.38  pounds  of  copper.  A  convenient  way 
of  making  up  the  electrolyte  is  to  take  by  weight: 
water,  75  parts;  bluestone  (CuS04 +"5H20),  16  to 
18%.  After  the  bluestone  is  dissolved,  add  sulphuric 
acid  to  make  up  to  16  to  18°  Beaume  (sp.  gr.  1.12 
to  1.15).  A  small  amount  of  salt  is  sometimes  added 
to  precipitate  antimony  in  the  electrolyte  as  oxy- 
chloride  and  also  to  act  as  a  guard  and  prevent  the 
silver  from  going  into  solution. 

While  there  are  many  so-called  processes  of  elec- 
trolytic refining,  all  are  ultimately  reduced  to  two 
systems,  and  according  as  the  electrodes  in  such 
cases  are  in  multiple  arc  or  in  series  they  are  known 
as  the  multiple  and  the  series  systems,  respectively. 
In  the  multiple  system  an  arrangement  is  contem- 
plated where  the  electrodes  are  in  parallel  arc,  the 
tanks  being  in  series ;  while  in  the  series  system  there 
is  intended  an  arrangement  where  the  electrodes  are 


102  TESTING  FOR  PROCESSES 

in  series.  The  tanks  may  be  either  in  series  or  mul- 
tiple series.  From  a  consideration  of  electrical  prin- 
ciples it  is  obvious  that  the  voltage  required  per  tank 
will  depend,  neglecting  the  constant  of  the  solution, 
upon  the  area  of  the  plates,  their  spacing,  and  the 
number  in  series.  The  multiple  system  is  applicable 
to  all  grades  of  copper,  it  permits  the  use  of  both 
high  and  low  current-density  in  lead-lined  baths,  and 
it  can  be  carried  out  with  little  Joss  of  current  as 
compared  with  the  series  system. 

Procedure. — Cast  an  anode  of  copper  which  con- 
tains impurities  usually  found  in  anode  copper,  such 
as  O2,  S,  As,  Sb,  Fe,  Pb,  Au,  and  Ag.  Before  pour- 
ing, take  a  sample  of  the  molten  copper  and  gran- 
ulate by  pouring  slowly  from  the  ladle  into  a  pail 
of  water;  reserve  this  sample  and  assay  for  copper 
and  impurities.  Prepare  the  anode  mold  by  giving 
it  a  wash  of  clay,  lampblack,  or  graphite  and  heating 
on  top  of  the  furnace.  The  anode  should  be  cast 
approximately  one-fourth  of  an  inch  in  thickness. 
Prepare  electrolyte  to  fill  one  tank  and  provide  for 
circulation.  In  case  sufficient  electrolyte  is  left  from 
a  previous  experiment,  assay  for  CuSO4  and  H2SO4 
and  bring  up  to  the  required  strength.  Carefully 
weigh  the  anode  and  place  in  position  in  the  tank. 
Care  should  be  taken  to  have  all  contact  surfaces 
clean  and  bright.  They  should  be  flat  so  as  to  give 
as  little  resistance  as  possible.  Weigh  the  cathodes 
and  place  in  position,  using  one  more  cathode  than 
the  number  of  anodes  in  the  tank. 

When  the  anodes  and  cathodes  are  in  position  and 
the  circulation  of  the  electrolyte  started,  throw  on 


ELECTROLYTIC  PROCESSES  103 

the  current  from  the  dynamo,  using  15  amperes  per 
square  foot  of  anode  surface  exposed.  Analyze  the 
electrolyte  each  day  for  CuSO4  and  H2S04  and  keep 
up  to  the  required  standard  by  the  addition  of  blue- 
stone  or  sulphuric  acid  as  required.  Observe  all 
precautions  mentioned  in  the  notes  so  as  to  obtain 
a  firm,  pure  deposit  of  copper  on  the  cathode.  Keep 
a  log  showing  each  day  the  number  of  ampere  hours 
run,  the  number  of  anodes  in  the  circuit  and  their 
arrangement,  analysis  of  the  electrolyte,  drop  in 
voltage  between  each  anode  and  cathode,  appear- 
ances of  the  deposit,  and  general  remarks.  The  drop 
in  voltage  between  the  plates  is  measured  by  a 
forked  stick  provided  with  contact  wires  leading  to 
a  volt-meter.  This  precaution  is  taken  to  detect 
any  short-circuit  and  observe  any  unduly  high  re- 
sistance from  poor  contacts  or  from  coatings  of 
slime  on  the  anodes.  The  latter  may  be  removed  by 
scrubbing  with  a  stiff  brush. 

At  some  period  during  the  run  prepare  at  least 
two  stripping  plates  as  follows:  Take  a  regular 
anode  made  of  copper  sheeting  from  %  to  1/16  in. 
thick,  make  a  coating  of  paraffine  %  m-  wide  around 
the  edges,  covering  the  remainder  of  the  surface  with 
graphite.  Hang  in  the  circuit  in  place  of  one  of 
the  regular  cathodes  and  allow  the  copper  to  de- 
posit for  24  hours.  The  paraffine  on  the  edges  of 
the  plate  and  the  graphite  coating  allow  the  result- 
ing deposit  to  be  easily  ' stripped'  off  in  a  sheet 
which  is  used  as  a  cathode  in  succeeding  runs. 

At  the  end  of  the  run  the  anodes  and  cathodes 
are  removed,  washed,  dried,  and  carefully  weighed. 


104  TESTING  FOR  PROCESSES 

Save  the  slime  washed  from  the  anode  and  add  to 
that  taken  from  the  bottom  of  the  tank.  The  loss 
in  weight  of  the  anodes  should  approximately  equal 
the  increase  in  weight  of  the  cathodes  (do  not  neg- 
lect copper  deposited  on  stripping  plates)  plus  the 
weight  of  the  slime.  Compute  the  efficiency  of  the 
run.  Knowing  the  number  of  ampere  hours  run, 
the  theoretical  deposition  can  then  be  determined 
as  follows: 

actual  deposition 

X  100  =  per  cent  efficiency 

theoretical  deposition 

Make  an  analysis  of  the  cathode  copper  for  ar- 
senic and  other  impurities.  The  percentages  of  these 
present  will  indicate  the  efficiency  of  the  run.  Care- 
fully siphon  off  the  electrolyte  from  the  tank  so  as 
not  to  disturb  the  sediment.  Recover  the  slime  by 
draining  the  contents  of  the  bottom  of  the  tank 
through  a  large  filter.  Wash  free  from  acid  and 
salts,  dry,  and  refine.  Recover  the  gold  and  silver 
from  the  bullion  by  cupellation  and  parting. 

ELECTROLYTIC  REFINING   OF  LEAD— BETTS 

PROCESS 

Principles/ — The  process  is  based  on  the  fact  that 
lead  is  easily  soluble  in  an  acid  solution  of  lead 
fluosilicate,  which  possesses  both  stability  under  elec- 
trolysis and  high  conductivity,  and  from  which  ex- 
ceptionally pure  lead  may  be  deposited  from  impure 
anodes.  The  cost  of  the  operation  is  comparatively 
low.  With  such  a  solution  there  is  no  polarization 
from  the  formation  of  lead  peroxide  at  the  anode, 
no  evaporation  of  the  constituents,  except  water, 


ELECTROLYTIC  PROCESSES  105 

and  no  danger  in  handling.  In  order  to  obtain  a 
firm  deposit  of  lead  on  the  cathode  it  has  been  found 
necessary  to  make  the  solution  reducing  by  the  addi- 
tion of  gelatine  or  pyrogallol.  During  the  electrol- 
ysis the  SiF6  ions  travel  toward  the  anodes  and  there 
combine  with  the  lead.  The  lead  and  hydrogen 
travel  in  the  opposite  direction  and  out  of  the  slime. 
There  are  comparatively  few  lead  ions  present,  so 
that  the  solution  in  the  neighborhood  of  the  anodes 
must  increase  in  concentration  and  tend  to  become 
neutral.  This  causes  an  E.  M.  F.  of  polarization 
to  act  against  the  E.  M.  F.  of  the  dynamo,  amounting 
to  approximately  0.2  of  a  volt  for  each  tank.  The 
greater  effect  comes  from  the  neutral  solution  with 
which  the  slime  is  saturated.  There  is,  consequently, 
an  advantage  in  working  with  rather  thin  anodes 
when  the  bullion  is  impure  enough  to  leave  the  slime 
sticking  to  the  plates.  The  fact  that  the  slime  sticks 
to  the  anode  is  compensated  by  the  increased  ease 
of  handling  the  slime,  by  removing  with  the  anodes 
to  special  cleaning  and  washing  tanks. 

Lead  stands  higher  in  the  scale  of  electromotive 
forces  of  solution  than  any  of  the  impurities  that 
it  contains  in  appreciable  amounts,  so  that  the  elec- 
trolyte does  not  need  to  be  changed  or  purified,  as 
in  copper  refining,  when  treating  the  ordinary 
grades  of  bullion.  The  electro-chemical  equivalent 
of  lead  is  also  high ;  one  ampere  depositing  3.88 
grams  of  lead  per  hour,  or  3^  as  much  lead  as  cop- 
per in  the  ordinary  copper-refining  solution. 

The  metallic  elements  that  enter  into  considera- 
tion as  possible  constituents  of  the  electrolyte  are 


106  TESTING  FOR  PROCESSES 

the  elements  usually  present  in  the  lead  bullion, 
those  that  may  be  in  the  fluosilicic  acid  as  impurities 
at  the  start,  and  iron  from  any  exposed  binding  of 
the  tanks.  Arsenic,  antimony,  seleninum,  and  tellu- 
rium are  easily  precipitated  by  lead,  and  conse- 
quently if  they  get  into  solution  they  will  be  thrown 
down  by  the  lead  electrodes;  probably  by  the  cath- 
odes, for  the  anodes  are  usually  covered  with  slime. 
This  would  prevent  their  reducing,  for  instance, 
much  antimony,  although  the  antimony  of  the  slime 
would  throw  out  such  an  easily  precipitated  metal 
as  silver.  Zinc,  iron,  and  nickel,  if  they  find  their 
way  into  the  solution,  stay  there;  for  they  are  not 
precipitated  by  the  lead  electrodes,  nor  can  they  be 
in  any  way  thrown  out  on  the  cathodes  by  the  elec- 
tric current  so  long  as  there  is  a  fair  amount  of  lead 
in  solution,  as  is  always  true.  It  is  doubtful  whether 
the  iron  goes  into  solution  at  all,  as  the  slime  usually 
assays  from  %  to  2%  of  this  metal.  Tin  occupies 
practically  the  same  position  in  the  scale  of  electro- 
motive forces  that  lead  does,  consequently  a  mixture 
of  lead  and  tin  will  behave  practically  as  one  metal. 
Alloys  of  copper  and  lead  give  a  slime  practically 
free  from  lead.  Alloys  of  40%  copper  and  60  lead, 
can  be  successfully  treated.  Silver  retains  very 
little  lead  in  the  slime.  Bismuth  holds  back  1/6  and 
antimony  Y4  to  1/5  its  weight  in  lead. 

The  electrolyte  is  prepared  by  saturating  35% 
hydrofluoric  acid  with  quartz,  adding  the  required 
amount  of  lead  as  carbonate  and  gelatine  in  the  form 
of  glue.  The  amount  of  gelatine  consumed  under 
good  working  conditions  is  only  %  to  %  pound  to 


ELECTROLYTIC  PROCESSES  101 

the  ton  of  lead  deposited.  Gelatine  in  the  form  of 
glue  is  always  used  and  is  cheaper.  The  better  grade 
of  glue  should  be  used,  as  the  cheaper  varieties  make 
a  disagreeable  smell  in  the  tank  room.  The  lead 
deposit,  forming  in  lead-fluosilicate  fluosilicic  acid 
solution,  containing  0.1%  gelatine  and  5%  or  more 
of  lead,  is  smooth  and  solid,  and  pieces  cut  from 
the  cathode  show  a  specific  gravity  of  11.35  to  11.40 
(the  specific  gravity  of  cast  lead).  With  a  little 
more  lead  and  the  average  current  density  employed 
in  commercial  operations,  about  15  amperes  per 
square  foot,  the  resulting  cathodes,  after  reaching 
a  considerable  thickness,  are  smoother.  The  smooth- 
ness and  purity  of  the  deposited  lead  are  propor- 
tional. Most  of  the  impurities  seem  to  be  introduced 
mechanically  through  the  attachment  of  floating  par- 
ticles of  slime  to  the  irregularities  of  the  cathode. 

The  anodes  and  cathodes  are  spaced  from  1%  to 
2  inches.  A  current  strength  of  10  to  25  amperes 
has  been  used,  but  14  amperes  gives  the  best  results 
as  regards  the  economy  of  working  and  the  physical 
and  chemical  properties  of  the  refined  metal  pro- 
duced. 

Procedure. — The  method  in  treating  lead  is  prac- 
tically the  same  as  in  the  electrolytic  refining  of 
copper.  The  fluosilicic  acid  is  prepared  by  putting 
hydrofluoric  acid  of  15  to  20%  strength  in  a  lead 
pan  and  adding  an  excess  of  powdered  quartz  or, 
better,  calcined  flint  which  dissolves  more  readily. 
Heat  the  pan,  but  not  so  much  as  to  cause  the  acid 
to  boil,  until  the  solution  is  saturated  with  silica, 
or  until  the  pungent  odor  of  HF  has  ceased.  To 


108'  TESTING  FOR  PROCESSES 

make  the  lead  solution,  add  the  required  amount 
of  white   lead,   which   ordinarily   contains   80%    of 
metallic  lead.    Gelatine  or  glue  is  added  to  the  elec- 
trolyte in  the  form  of  a  strong  hot  solution  in  water. 
The  electrolyte  is  made  up  so  as  to  contain : 
10     to  12%  H2SiF6 
5      "      9%  Pb 
0.1  "      0.2%  gelatine 
Sp.  gr.  1.16  (approximately) 
Temperature  during  electrolysis  34°  C. 
To  determine  acidity  of  electrolyte,  add  alcoholic 
potassium  acetate  solution.    Filter  off  K2SiF6,  wash 
with  diluted  alcohol,   add  filter  and  precipitate  to 
distilled  water  in  a  beaker,  heat  to  boiling  and  titrate 
with  NaOH,*  using  rosolic  acid  preferably,  but  also 
phenolphtalein  as  an  indicator.     To  determine  lead, 
add    H2S04,    filter,    and    determine    by    molybdate 
method.      Other   determinations    are    seldom   made. 
For  more  detailed  information,  and  for  methods  of 
slime  treatment  see  'Lead  Refining  by  Electrolysis' 
by  Betts. 

*1  gram  of  NaOH  =  0.9  gram  H2SiF6. 


ZINC  SMELTING 

General  Process. — The  zinc-bearing  minerals  sel- 
dom occur  pure  and  are  usually  associated  with  other 
metaliferous,  silicious  or  earthy  gangue  minerals. 
Chemically  combined  impurities  such  as  iron  are 
also  almost  invariably  present.  The  mechanically 
combined  minerals  may  be  separated  to  a  greater 
or  less  extent  by  methods  of  ore  dressing,  but  for 
the  elimination  of  chemically  combined  impurities 
a  metallurgical  treatment  is  required. 

The  carbonates  and  silicates  of  zinc  are  usually 
calcined  preliminary  to  retorting,  in  order  to  drive 
off  the  carbon-dioxide  and  water.  In  roasting  blende 
preliminary  to  the  distillation  of  the  zinc,  the  aim 
is  to  convert  the  sulphide  as  completely  as  possible 
into  the  oxide,  since  every  part  of  sulphur  remain- 
ing behind  in  the  roasted  ore  means  approximately 
two  parts  of  zinc  held  in  the  retort  after  distillation. 
The  accomplishment  of  this  object  requires  fine  corn- 
munition  of  the  zinc  blende,  generally  not  to  exceed 
2  mm.  size  (6-mesh),  high  temperature,  and  slow, 
careful  roasting  with  frequent  stirring.  It  is  never 
economical  to  effect  a  complete  elimination  of  the 
sulphur,  which  is  impossible  with  some  ores.  In 
good  practice  the  amount  in  combination  with  zinc, 
iron,  and  lead  is  reduced  to  one  per  cent  and  fre- 
quently lower.  Ores  containing  lime  and  magnesia 
may  retain  much  more  owing  to  the  formation  of 
calcium  and  magnesium  sulphates.  These  last  two 
sulphates  are  formed  by  the  action  of  sulphuric  an- 


110  TESTING  FOR  PROCESSES 

hydride  on  their  respective  oxides.  Calcium  sulphate 
is  not  decomposed  at  the  temperature  attained  in 
the  roasting  furnace,  but  magnesium  sulphate  is 
partly  decomposed  in  the  last  stages  of  the  process. 

Chemical  Reactions. — Starting  at  a  dull  red  heat, 
ZnS  +  30  =  ZnO  +  S02 

-  41880  =  +  85430  +  71000  =  +  114550  calories 

This  reaction  is  exothermic  and  develops  sufficient 
heat  to  proceed  independently  until  the  sulphur  is 
burned  down  to  5  to  8%  according  to  the  conditions. 

At  a  bright  red  heat,  S02  +  0  =  S03,  by  contact 
(catalysis)  with  the  glowing  ore  and  furnace  walls, 
and  ZnO  +  S03  =  ZnS04. 

At  a  cherry-red  heat  the  zinc  sulphate  breaks  up 
into  basic  sulphate  (3ZnO.ZnOS04)  and  S03.  The 
S03  is  in  turn  broken  up  more  or  less  into  S02  and 
0.  The  basic  zinc  sulphate  is  not  broken  up  entirely 
until  it  has  been  exposed  to  a  bright  red  heat  for  a 
considerable  time  where  there  is  danger  of  the  loss 
of  zinc  by  volatilization,  and  the  sintering  of  the 
charge  when  lead,  iron,  or  manganese  are  present. 
An  experienced  roaster  can  judge  by  the  volume 
and  cessation  of  the  fumes  over  the  ore  the  different 
stages  and  time  of  completion  of  the  roast.  The 
presence  of  sulphates,  which  is  an  important  factor, 
can  only  be  determined  by  analysis. 

Distillation. — The  zinc  in  the  ore,  having  been  con- 
verted into  oxide  by  calcinating  or  roasting,  the  re- 
covery of  the  metal  is  accomplished  by  virtue  of  the 
fact  that  the  oxide  is  reducible,  when  heated  to  a 
high  temperature,  by  carbon  or  carbon  monoxide  or 
both,  according  to  the  reaction: 


ZINC  SMELTING  111 


-  86000  =  +  29000  =  -  57000  calories 
This  reaction  takes  place,  according  to  different 
authorities,  from  800°  C.  to  1075°  C.,  but  practically 
a  temperature  of  1300°  C.  is  required.  The  metal 
is  reduced  in  the  form  of  vapor  which  is  subse- 
quently liquefied  by  cooling  to  about  415  to  550°  C. 
All  the  zinc  vapor  does  not  condense  as  a  liquid 
but  5  to  10%  of  it  passes  directly  into  the  solid  state 
in  the  form  of  a  bluish  powder  called  zinc  fume. 
The  presence  of  silica,  alumina,  lime,  magnesium, 
iron,  manganese,  and  lead  is  detrimental  when  they 
are  in  proportion  to  form  fusible  slags  which  corrode 
the  retort  and  coat  over  the  particles  of  ZnO,  thus 
causing  a  loss  of  zinc  in  the  residues.  When  cad- 
mium, arsenic,  and  antimony  are  present  they  are 
partly  condensed  with  the  zinc. 

Procedure  in  Testing.  —  Take  a  500-gram  sample 
of  the  ore  and  crush  to  8  mesh.  Sample  and  assay 
the  original  ore  for  zinc,  sulphur,  and  impurities. 
Place  the  500  grams  of  ore  in  a  large  roasting  dish 
and  put  in  place  on  the  hearth  of  the  gas-oven  fur- 
nace. Observing  carefully  all  the  precautions  pre- 
viously given,  roast  down  to  the  lowest  practical 
limit  of  sulphur.  At  the  end  of  the  operation  weigh 
the  ore  and  take  a  sample  and  assay  for  sulphur 
and  zinc.  Mix  the  roast  ore  with  50%  of  its  weight 
of  crushed  coal  and  charge  into  a  retort,  which  has 
been  slowly  brought  up  to  a  white  heat  in  the  gas 
forge.  Put  the  condenser  in  place  and  keep  at  a 
good  white  heat  until  the  fumes  of  zinc  cease  to  ap- 
pear at  the  mouth  of  the  condenser.  Allow  to  cool 


112  TESTING  FOR  PROCESSES 

down  slowly,  remove  the  residue  from  the  retort, 
weigh,  sample,  and  assay  for  zinc.  Collect  the  spel- 
ter and  zinc  fume  and  weigh  separately  and  assay 
for  zinc.  Note  the  condition  of  the  interior  of  the 
retort. 

Report. — Efficiency  of  the  roast,  that  is,  loss  of 
zinc  and  sulphur,  percentage  of  recovery  of  zinc  by 
retorting,  cost  data,  should  be  tabulated  for  ready 
reference. 


FURNACE  TEST 

General. — The  object  of  this  test  is  to  determine 
as  nearly  as  possible  the  heat  efficiency  of  a  furnace 
as  well  as  the  losses.  The  starting  point  will  evi- 
dently be  the  determination  of  the  calorific  power  of 
the  fuel  used.  This  will  furnish  means  for  a  com- 
parison of  the  quantities  of  heat  lost  and  utilized  in 
the  furnace.  The  next  point  is  to  construct  a  heat 
balance  sheet  of  the  furnace  in  question  for  the  pur- 
pose of  placing  in  a.  clear  and  concise  manner  the 
ratio  of  heat  used  to  the  total  available  heat,  also 
to  the  heat  losses.  The  value  of  a  furnace  can  be 
determined  with  a  certainty,  and  if  the  efficiency 
is  below  a  required  standard  the  cause  of  such  a 
deficiency  may  be  deducted  from  the  test. 

The  determination  of  the  heat  balance  requires 
the  following:  (1)  heat  utilized;  (2)  heat  carried 
away  by  the  waste  products;  (3)  heat  lost  by  con- 
duction and  radiation.  The  sum  of  these  three  quan- 
tities must  be  equal  to  the  total  available  heat,  the 
calorific  power  of  the  fuel.  It  also  follows  that  when 
any  two  of  the  quantities  are  known,  the  third  may 
be  determined  by  difference. 

The  experimental  data  required  are  the  follow- 
ing: (1)  calorific  power  of  the  fuel;  (2)  elementary 
analysis  of  the  fuel  and  ash;  (3)  analysis  of  the 
gaseous  products;  (4)  temperature  determinations 
at  various  points  in  the  furnace. 

Before  proceeding  with  the  test,  determine  the 
specific  object  of  the  proposed  trial,  whether  it  is 


114  TESTING  FOR  PROCESSES 

to  ascertain  the  efficiency  of  the  furnace  or  its  de- 
fects, the  economy  of  a  particular  kind  of  fuel,  or 
the  effects  of  change  of  design,  proportion  or  opera- 
tion, and  prepare  for  the  trial  accordingly.  Ascer- 
tain the  dimensions  of  the  grates  and  important 
parts.  Clean  the  furnace  thoroughly  and  stop  any 
air  leaks  in  the  setting.  Before  the  trial  see  that 
the  furnace  is  heated  to  its  usual  working  tempera- 
ture. If  it  has  been  laid  off  and  become  cold,  it 
should  be  worked  before  the  trial  until  the  walls 
are  well  heated. 

The  standard  method  of  starting  and  stopping  a 
test  is  with  clean  grate-bars  and  ash-pit,  but  the 
alternate  method,  while  not  so  accurate,  is  more  con- 
venient. The  fire  is  burned  low  and  well  cleaned, 
and  all  the  ashes  removed.  Note  the  amount  of  coal 
left  on  the  grate,  as  nearly  as  can  be  estimated  by 
the  eye,  and  take  the  temperature  of  the  flue  gases. 
At  the  end  of  the  test  the  fires  should  be  burned  low 
and  cleaned  so  as  to  have  the  conditions  as  nearly 
as  possible  the  same  as  at  the  start.  Fair  samples 
of  the  coal  and  ash  are  to  be  taken  and  enclosed  in 
air-tight  cans.  The  total  weight  of  both  coal  and  ash 
are  to  be  accurately  determined.  A  sample  of  the 
flue  gas  is  drawn  off  continuously  during  the  experi- 
ment by  means  of  an  aspirator.  The  temperatures 
of  the  flue  gases  are  taken  by  means  of  a  thermo- 
electric pyrometer.  The  hot  junction  is  placed  near 
the  gas-sampling  tube  and  the  lead  wires  run  to  a 
convenient  place  for  the  galvanometer.  The  stack 
suction  is  determined  by  means  of  a  U  tube,  or  pref- 
erably with  a  draft  gauge  graduated  to  read  to  1/100 


FURNACE  TEST  115 

of  an  inch  water  column.  A  quarter-inch  iron  pipe  is 
inserted  in  the  flue  and  connected  to  the  draft  gauge 
by  means  of  a  suitable  length  of  rubber  tubing. 

Resume. — Put  all  the  apparatus  in  place  for  tak- 
ing the  necessary  reading. 

Introduce  into  a  hole  in  the  stack  the  gas-sampling 
tube,  pyrometer  junction,  and  the  draught-gauge 
tube  as  shown  in  Pig.  16.  Place  the  instruments 
where  they  will  not  be  affected  by  the  heat,  not  liable 
to  damage  and  can  be  easily  read. 

Clean  the  fire  and  ash-pit.  Note  the  thickness  of 
fire  and  condition  of  the  bed,  also  temperature  of 
flue  gas. 

Weigh  out  enough  coal  for  the  first  firing,  as 
nearly  as  can  be  judged ;  take  a  sample  of  the  coal. 
Take  initial  readings  of  the  galvanometer  and  draft 
gauge  and  start  the  aspirator  to  taking  the  gas 
sample. 

Put  coal  on  the  dead  plate  and  a  thin  layer  over 
the  bed  of  the  fire.  Note  the  time  of  commencing 
and  of  firing. 

Take  necessary  readings  of  draft  gauge  and  py- 
rometer every  five  minutes. 

Fire  up,  when  necessary,  to  keep  the  muffles  at  a 
temperature  required  for  good  assay  work. 

Run  the  gas  aspirator  so  as  to  fill  two  sample 
bulbs  during  the  hours  run.  Let  the  gas  be  drawn 
through  the  sample  tube  for  ten  minutes  before  re- 
moving it,  so  as  to  have  the  flue  gas  thoroughly  re- 
place the  air  in  the  tube. 

At  the  end  of  the  test  take  a  sample  of  the  ash 
and  determine  its  sensible  heat  by  dumping  into  a 


FURNACE  TEST 


117 


pail  containing  a  known  amount  of  water  and  weight. 

Make  an  analysis  of  the  two  gas  samples  with  the 
Orsat  gas  apparatus,  determining  C02,  CO,  0,  and 
N  (by  difference). 

Make  analyses  of  coal  and  ash,  and  determine  their 
calorific  power. 

Report 

Calorific  value  1  Ib.  coal  == Ib.  cal  100%. 


Distribution 

Lb.Coal 

% 

Heat  used  usefully  

Calorific  value  ash  from  1  Ib.  coal  

Loss  due  to  moisture  in  coal* 

Loss  due  to  sensible  heat  carried  away 
in  dry  chimney  gases  

Loss   due   to  incomplete  combustion   of 
carbon,  etc.,  in  flue  gases  

Loss  due  to  moisture  in  the  air,  radia- 
tion and  unaccounted  for,  to  make 
up  to  100% 

Totals  

*A11  calculations  computed  on  the  basis  of  1  Ib  coal. 

Pounds  of  coal  burned  per  square  foot  of  grate 
area,  per  hour  = 


METHODS  OF  CALCULATION  FOR  HEAT- 
BALANCE  SHEET 

Heat  Used  Usefully. — The  calculations  under  this 
head  will  vary  with  the  nature  of  the  work  done  in 
the  furnace  and  is  not  taken  into  account  in  the  test 
of  assay  furnaces.  In  roasting,  the  heat  equations 
of  the  various  constituents  of  the  ore  must  be  in- 
cluded. In  fusion,  the  heat  required  to  fuse  the 
charge  must  be  taken  into  account. 

Loss  of  Heat  Due  to  Moisture  in  the  Coal. — This 

may  be  determined  by  the  following  formula : 

P  =  per  cent  moisture  in  the  coal. 

t  =  temperature  of  the  room. 

T  =  temperature  of  the  flue  gas. 

C  =  thermal  capacity  of  water  vapor  at  tempera- 
ture t. 

d  =  thermal  capacity  of  water  vapor  at  tempera- 
ture T. 

^  X  ( (100  - 1)  +  518  +  (C,  -  C) )  =  calories. 

Where  100  =  boiling  point  of  water. 

518  =  latent  heat  of  vaporization  of  water. 

The  values  for  Cj  and  C  are  read  off  the  curve ; 
see  Fig.  17. 

Loss  Due  to  Heat  Carried  Away  in  Dry  Chimney 
Gases. — This  equals  the  sum  of  the  thermal  capaci- 
ties of  the  constituents  of  the  flue  gas. 

The  thermal  capacity  of  each  gas  is, 
weight  gas  X  (C1-C)=Q 


\ 


120  TESTING  FOR  PROCESSES 

The  combined  thermal  capacities  would  be, 

Wt.  C02X  Q  +  wt.  CO  X  Q  +  wt.  OX  Q  +  wt.  N  X  Q 
-f-  etc.  =  calories. 

To  find  the  weight  of  the  several  gases  proceed  as 
follows : 

Suppose  the  gas  to  contain  C02,  CO,  0,  and  N. 

(1)  The  molecular  volume  of  any  gas  is  22.38 
liters  =  M,  or  the  molecular  weight  of  any  gas  in 
grams  has  a  volume  of  22.38  liters  =  M. 

(2)  Let  WC02,  WCO,  etc.,  stand  for  the  molecu- 
lar weights  of  the  various  substances. 

(3)  Let  C02,  CO,  etc.,  stand  for  the  percentages 
by  volume. 

In  one  unit  volume  of  gas  we  have, 

^  ^100  +  100  =  volume  of  c°2  +  CO  in  1  unit 
volume  of  gas. 

(5)  According  to  (1)  and  (2)  if  WC02  =  M,  1 
unit  volume  of  C02  will  weigh  W^°?  and  similarly 

1  unit  volume   of  CO  will  weigh  ' 

But  in  (4)  we  have  the  volume  of  C02  +  CO. 

(a\     CQ2  v  WCO2    ,    CO         WCO 
w     100  X      M       h  100    X     M 
CO  in  1  unit  volume. 


(7)     But   C02   is  12/44   carbon  -     and   CO 

is  12/28  carbon. 

CO2  WCO2  WC  CO  WCO  WC 

100  M      X    WCOa   "      TOO   X       M       X  WCO 

=  weight  carbon  in  1  unit  volume. 


HEAT  BALANCE  121 

Combining  and  substituting  values, 

(8)  12      (C02  +  CO)  =  weight  carbon  in  1  unit 
2238 

volume. 

The  carbon  in  the  fuel  —  the  carbon  in  the  ash  = 
carbon  burned  to  flue  gas. 

From  (8)  we  have  the  weight  of  carbon  in  1  vol- 
ume of  flue  gas. 

From  1  unit  weight  of  carbon  burned  to  flue  gas 
containing  C02,  CO,  0,  N,  we  will  get, 

l 

(9)  12(CO2  +  CO)     =  unit  volumes  of  flue  gas. 
2238 

According  to  (1)  —  (4)  and  (5), 
(10)     Weight  C02  in  1  volume  gas 


CO 
0 

N 


%  gas  X  W 
100M 


And  from  (9)  the  unit  volumes  of  gas. 
Then, 

C03  +  WCO,  1  wt.C02froml 

(11)  100M         ^12(C02  +  CQ)  =unit   of  carbon 

223§  burned. 

Substituting  values  and  combining. 

(12)  11CO2        __  weight  of  CO3  from  1  unit  of  car- 
3(C03  +  CO)  ~bon  burned. 

In  a  like  manner, 

7CO  weight  of  CO  from  1  unit  of  car- 


(13) 


3(COa 


=  bon  burned. 


122  TESTING  FOR  PROCESSES 

7N  __  weight  of  O  from  1  unit  of  carbon 
3  (CO,  -f  CO)  ~~  burned. 

8Q          __  weight  of  N  from  1  unit  of  carbon 
3(CO2  +  CO)  ~~  burned. 

Now  having  the  weight  of  the  constituent  gases 
per  unit  weight  of-  carbon  burned,  the  same  per 
pound  of  coal  burned  may  be  calculated. 

a  =  weight  of  coal  burned. 

b  =  weight  of  ash  obtained. 

c  =  percentages  of  carbon  in  coal. 

d==  "  "        "        "    ash. 


100  = 


ac  —  bd 
Or,  -  -  -  —  per  cent  available  carbon  in  the  coal. 

Therefore  the  weight  of  any  constituent  gas  X  per 
cent  available  carbon  in  the  coal  -f-  100  =  weight  of 
that  gas  per  unit  weight  of  coal  burned. 

80  ac  —  bd 

(17)  Example.  3(CQ3  +  CO)        looa"     -weight 

of  0  per  pound  of  coal. 

As  previously  given  the  heat  carried  away  by  a 
gas  per  unit  of  weight  of  coal. 

(18)  Wt.  gas  per  Ib.  coal  (d  -  C)  =  calories. 

d  =  thermal  capacity  of  gas  at  stack  temperature. 

C  =  "         "    "    "  room        " 

The  value  for  C  and  d  are  taken  from  the  curve, 

Fig.  17  or  may  be  computed  for  any  temperature 

as  follows: 

(19)  Q  =  aXt  +  bXlO-  6t2. 


HEAT  BALANCE  123 

Where  a  and  b  are  factors  variable  with  different 
gases. 

a  b 

Oxygen,  O 0.213  19 

Nitrogen,  N;  carbon  monoxide,  CO 0.234  21 

Hydrogen,  H  0.340  300 

Water  vapor,  H20 0.447  162 

Carbon  dioxide,  COa 0.193  84 

Methane,  CH4 0.608  374 

Total  heat  carried  away  in  dry  chimney  gases, 
Let  w  =  weight  of  gas  per  pound  of  coal  (17) 
Q  =  thermal  capacity. 
(20)    wC02Q  +  wCOQ  +  wOQ  +  wNQ  =  calories. 

(d)  Loss  due  to  incomplete  combustion  of  carbon 
in  the  flue  gases,  represented  by  the  available  calories 
of  the  CO. 

As  given  under  Ex.  (17) 

/o-i  \  7CQ  v   ac— *>d  =    weight  of  CO  per 

3(CO2  +  CO)  100  a          pound  of  coal. 

Calorific  value  of  CO  =  2436. 
(22)    Weight  of  CO  per  pound  of  coal  X  2436  = 
calories  available  in  CO. 

(e)  Heat  value  of  ash.     See  ' Calorimetry. '     (P. 
123.) 

Heat  value  of  1  pound  of  ash  times  weight  of  ash 
obtained  divided  by  weight  of  coal  burned  equals 
the  heat  value  of  ash  per  pound  of  coal. 

(f)  Sensible    heat    of   ash.     See    'Furnace    Test, 
Resume.'     (P.  113.) 

In  the  above  calculations  ethylene,  C2H,  methane, 
CH4,  etc.,  are  not  taken  into  account.  If  separate 
determinations  of  these  gases  are  obtained  the  re- 


124  TESTING  FOR  PROCESSES 

suits  may  be  used  to  make  the  balance  sheet  more 
complete  and  accurate,  by  making  corrections  for 
these  gases  along  lines  already  given. 
For  instance  (8)  would  be  changed  to, 

1  unit    volumes    of    flue 

(23)  12  (CQ3  +  CO  +  CaH4  -f  CH4)  -  gas  per  unit  weight  of 
2238  2  carbon  burned. 

Letting  3  (COa  +  CO  +  ^Mi    ,    CH4)    =  a. 
2 

(24)  11CQa  ==  weight  COa  per  unit  weight  carbon  burned. 

(25)  jgo=   «    co   •«    • 

(26)  -^-  =      "          N       ' 

a 

(27)  8Q   =      "          o 

a 

7r 

(28)  a 

=      "       CH4 


(30)     -JL=      »          H       " 

The  above  values  calculated  in  terms  of  a  unit 
weight  of  coal  burned  as  in  (17)  would  then  be  sub- 
stituted in  (20). 

QwC02  +  QwCO  +  QwO  +  QwN  +  QwC2H4,  etc. 
=  calories  carried  away  in  dry  gases. 


APPROXIMATE  DETERMINATION  OF  THE 
CALORIFIC  POWER  OF  A  FUEL 

Few  industrial  laboratories  are  equipped  with  cal- 
orimeters, and  it  is  often  desirable  to  find  within 
1%  (approximately)  the  calorific  value  of  a  fuel.  In 
this  case  the  Malher  and  Goutal  formulas  are  satis- 
factory and  have  stood  the  test  of  practice.  The 
Goutal  formula  makes  use  of  the  data  obtained  from 
the  ordinary  approximate  analysis ;  the  Malher  of 
the  ultimate  analysis. 

PROXIMATE  ANALYSIS 

Moisture  Determination. — The  moisture  is  deter- 
mined on  a  2-gram  sample  heated  for  one  hour  in  an 
air  bath  at  115 °C.  The  constant  weight  should  be 
verified  by  two  weighings. 

Volatile  Matter. — In  a  crucible  of  30  c.c.  capacity 
is  placed  a  5-gram  sample  of  the  powdered  fuel,  the 
crucible  is  placed  uncovered  over  a  Berzelius  burner. 
The  flame  is  maintained  so  as  to  completely  surround 
the  crucible.  After  all  illuminants  have  been  en- 
tirely driven  off,  heating  should  be  continued  for 
three  more  minutes. 

Ash. — Two  grams  of  fuel  are  burned  in  a  muffle 
until  only  a  white  residue  remains,  the  heating,  to 
be  done  very  carefully  and  not  too  suddenly. 

Fixed  Carbon. — The  fixed  carbon  is  determined  by 
difference  between  the  sum  of  the  other  constituents 
and  100  per  cent. 


126  TESTING  FOR  PROCESSES 

Example. — The  analysis  of  an  anthracite  is  as  fol- 
lows: 

Per  cent. 

Fixed  carbon  (C) 86.70 

Volatile  matter  (V) 10.05 

Ash   1.45 

Moisture   1.80 

100.00 

The  percentage  of  volatile  matter  V,  calculated  as 
if  there  were  neither  ash  nor  moisture  is, 
V  X  100  10.05  X  100 

V  :  C-f  V    =  :    86.70  +  10.05    =  10'04 

Goutal  Formula. — The  data  obtained  by  this  an- 
alysis may  be  used  in  calculating  the  approximate 
calorific  value  of  the  fuel  by  the  following  formula : 

C  =  fixed  carbon. 

V  =  volatile  matter. 

a  =  a  variable  factor. 

a  is  a  function  of  the  volatile  matter  in  the  fuel, 
allowance  being  made  for  moisture  and  ash. 

Calorific  value  =  82C  +  aV. 

For  all  fuels  for  volatile  matter  V  below  40%  the 
values  from  a  are  given  in  table  following. 

CALORIFIC  POWEB  OF  FUEL — GOUTAL  FACTORS 


V'%. 
1  to  5. 

a  Calories. 
100 

V'%. 
17  

a  Calories. 
113 

V'%.   a  Calories. 
29  99 

5  

145 

18 

112 

30 

98 

6  

142 

19  .. 

110 

31 

97 

7  

139 

20  ...  . 

.  .  .  109 

32 

97 

8  

136 

21  

.  .  .  108 

33 

96 

9  

133 

22  

107 

34  . 

95 

10  . 

.  130 

23  . 

.  105 

35  . 

94 

CALORIMETRY  OP  FUEL  127 


11  

127 

24  

104 

36  

91 

12  

124 

25  

103 

37  

88 

13 

122 

26 

102 

38 

85 

14 

120 

27  

.  .  .  .  101 

39 

82 

15  . 

117 

28  

...  100 

40  .. 

80 

16  . 

..115 

From  the  table  we  find  that  V'  (10.04)  corresponds 
to  a  value  of  a  equal  to  129  calories.  Substituting  in 
the  formula  we  have 

Calorific    power  =  82  X  86.7  +  129  X  10.05  =  8406 

calories. 

By  actual  determination  on  a  bomb  calorimeter  it 
was  found  to  be  8404  calories.  The  results  obtained 
with  the  formula  are  generally  within  1%  of  the  ex- 
perimental determinations,  and  an  error  of  2%  is 
quite  exceptional  and  only  observed  with  anthracites 
and  lignites.  For  the  more  exact  determinations  of 
the  calorific  power  of  a  fuel,  a  colorimeter  is  used. 
The  use  of  these  instruments  is  fully  explained  by 
booklets  furnished  with  the  outfit. 

MALHER  FORMULA 

When  an  ultimate  analysis  is  available  the  calorific 
power  may  be  calculated  as  follows : 
Q  =  calorific  power. 

8140C  +  34500H  —  3000  (O  +  N) . 
100 

Substituting  in  the  equation  we  finally  get. 

Q  =  111.400 +  375H- 3000  in  calories  per  kilo- 
gram of  fuel. 

This  formula  does  not  give  accurate  results  in  the 
case  of  cannel  coal,  lignite,  peat,  and  wood. 


PYROMETRY 

General. — In  the  preceding  chapters  it  has  been 
shown  that  it  is  necessary  in  making  certain  tests  to 
determine  accurately  and  quickly  the  comparatively 
high  temperatures  of  furnaces,  flues,  etc.  The  in- 
strument selected  for  the  work  should  be  portable, 
accurate,  easy  to  manipulate,  and  should  allow  the 
reading  of  temperatures  of  widely  distant  points. 
Of  the  many  types  the  LeChatelier  thermo-electric 
pyrometer  fulfills  best  the  above  mentioned  condi- 
tions. The  thermo-electric  instrument  being  small 
can  be  introduced  into  any  part  of  the  furnaces,  and 
since  a  few  seconds  are  sufficient  for  establishing  an 
equilibrium  of  temperatures,  it  serves  to  measure 
temperatures  close  on  to  1780°C.,  the  melting  point 
of  platinum,  without  endangering  the  life  of  the 
wires.  The  LeChatelier  pyrometer  is  thoroughly  re- 
liable, providing  care  is  taken  with  the  wires.  How- 
ever, in  order  to  be  successful  with  the  instrument, 
its  limitations  should  be  borne  in  mind.  These  are 
due  entirely  to  the  nature  of  the  metals  forming  the 
couple.  Practically  the  same  care  that  is  to  be  taken 
with  platinum  ware,  is  required  with  the  wires  of 
the  thermo-couple.  Platinum  is  readily  attacked  by 
the  vapors  of  volatile  metals.  Silver,  zinc,  antimony, 
and  copper  are  especially  injurious.  In  a  reducing 
atmosphere  silicates  and  phosphates  are  objection- 
able. In  spite  of  the  most  thorough  protection,  plati- 
num and  its  alloys  suffer  physical  alterations,  result- 
ing in  great  brittleness  by  mere  continuous  heating 
to  temperatures  above  1000 °C. 


PYROMETRY  129 

For  measuring  the  E.  M.  F.  of  the  thermo-electric 
couple,  a  galvonemeter  reading  to  millivolts  is  used. 
Any  two  metals  may  be  used  for  the  thermo-electric 
couple,  but  the  following  are  commonly  employed. 

1.  Pure  platinum  and  an  alloy  of  10%  rhodium 
and  90%  platinum  (or  10%  iridium  and  90%  plati- 
num). 

2.  Pure   silver    (or   copper)    and   alloy   of   25% 
nickel  and  75%  copper  (Constantan). 

3.  Pure  silver  and  an  alloy  of  25%  platinum  and 
75%  silver. 

4.  Pure  iridium  and  an  alloy  of  10%  ruthenium 
and  90%  iridium. 

The  first  alloy  is  used  for  all  temperatures  up  to 
1600 °C.  The  Constantan  couples  are  useful  for  low 
temperatures  between  zero  and  300°C.  They  have  a 
high  E.  M.  F.  with  the  further  advantage  of  being 
cheap.  The  silver  and  silver  platinum  couples  are 
good  for  all  temperatures  up  to  the  melting  point  of 
silver  (926°C).  This  couple  costs  much  less  than 
the  platinum  and  has  nearly  double  the  E.  M.  F.  The 
ruthenium-iridium  couple  is  useful  only  for  high 
temperature  measurements,  those  above  the  melting 
point  of  platinum.  It  is  extremely  brittle,  but  can 
be  used  up  to  2100°C. 

Principle  of  the  Thermo-Electric  Pyrometer. — 
Whenever  two  metals  in  contact  are  heated,  an  elec- 
tric current  is  generated  which  is  a  function  of  the 
temperature.  In  a  circuit  composed  of  a  number  of 
metals,  the  resulting  current  is  equal  to  the  algebraic 
sum  of  the  component  currents.  The  E.  M.  F.  of  a 
couple  is  not  altered  when  one  or  several  metals  are 


130 


TESTING  FOR  PROCESSES 


K 

tf 


inserted  between  the  metals  of  the  couples.     It  fol 
lows  that  any  metal  can  be  used  as  a  solder. 

Before  making  the  couples  the  wires  should  be  an 


PYROMETRY  131 

nealed  at  a  temperature  equal  to  or  even  higher 
than  that  to  which  they  will  be  exposed.  The  junc- 
tion should  be  made  by  twisting  the  wires  and  melt- 
ing the  twist  together  in  an  oxygen  blast  until  a 
globule  forms.  Care  should  be  taken  to  avoid  twist- 
ing and  kinking  the  wires  after  annealing.  Even 
in  perfectly  homogeneous  metals,  a  distorted  section 
will,  when  heated,  give  rise  to  an  electric  current 
which  will  of  course  affect  the  galvanometer  to  a 
greater  or  less  extent.  The  thermo-couple  should  al- 
ways be  soldered  to  the  copper  leads,  and  never  con- 
nected by  merely  twisting  them  together.  The  gal- 
vanometer may  be  placed  at  any  convenient  place,  a 
mile  removed  if  necessary.  Care  must  be  always 
taken,  in  order  to  insure  accuracy,  to  standardize 
the  couple  with  the  same  length  of  wires  as  that  used 
for  the  work.  The  necessary  length  of  copper  may 
actually  be  used  or  an  equivalent  resistance. 

It  is  to  be  remembered  that  the  thermo-electric 
current  generated  is  proportional,  not  to  the  actual 
temperature  of  the  thermo-j unction,  but  to  the  dif- 
ference in  temperature  of  the  hot  and  cold  junction. 
Care,  therefore,  should  be  taken  to  protect  the  con- 
nections of  the  thermo-junction  with  the  copper  wires 
(the  cold  junction)  from  undue  heating ;  in  case  this 
protection  is  not  possible  it  becomes  imperative  to 
use  a  water-cooled  end.  This  is  seldom  necessary  as 
it  is  usually  sufficient  to  take  the  temperature  of  the 
cold  junction  as  that  of  the  surrounding  atmosphere. 

^Standardization  of  a  Pyrometer. — In  order  to  find 

*Pyrometers  and  other  instruments  will  be  standardized 
for  small  fee  at  the  Bureau  of  Standards,  Washington,  D.  C. 


132  TESTING  FOR  PROCESSES 

a  temperature  corresponding  to  a  given  deflection  of 
the  galvanometer  in  millivolts,  it  is  necessary  to  find 
the  number  of  millivolts  corresponding  to  known  tem- 
peratures. Such  temperatures  are  given  by  the  melt- 
ing or  boiling  points  of  chemical  elements  or  definite 
chemical  compounds.  Mathematically  speaking,  with 
pure  platinum  and  iridio  and  rhodio-platinum 
couples,  it  is  sufficient  to  determine  two  points;  by 
means  of  a  logarithmic  formula  it  becomes  possible 
to  find  any  other  temperature  corresponding  to  a 
given  deflection.  However,  both  for  safety  and  ra- 
pidity it  is  better,  in  industrial  work,  to  use  the 
graphic  method.  If,  on  a  sheet  of  co-ordinate  paper, 
the  temperatures  are  plotted  on  one  axis  and  the  mil- 
livolts on  the  other,  the  intersections  of  the  ordinates 
and  abscissae  will  give  a  series  of  points  lying  on  a 
curve  characteristic  of  the  couple. 

The  fixed  points  to  be  used  are  as  follows: 

Degrees  C. 

Boiling  points  of  water,  H2O 100 

"    napthalene,  C10H8  218 

"          "       "    sulphur,  S    445 

Freezing  points  of  zinc,  Zn    419 

"  "       "    aluminum,  Al   655 

"    silver,  Ag  962 

"   copper,  Cu 1,084 

"    gold,  Au   1,065 

"    platinum,  Ft   1,780 

For  determining  the  deflections  for  the  boiling 
points  of  water,  napthalene,  and  sulphur,  use  the  ap- 
paratus shown  in  Fig.  19. 

The  bulb  test  tube  is  inserted  in  a  muff  of  plaster 
of  paris,  after  having  been  filled  to  about  one-third  of 


PYROMETRY 


133 


its  lower  length  with  a  substance  to  be  used;  heat 
is  carefully  applied  until  ebullition  has  begun.  Care 
should  be  taken  with  sulphur  and  napthalene  to  keep 
the  level  of  the  vapors  below  the  mouth  of  the  test 
tube,  otherwise  the  vapors  will  catch  fire.  For  these 


Fig.  19. 


-3-ir 


APPARATUS  FOR  THE  DETERMINATION  OF 
BOILING  POINTS 


determinations  it  is  not  usually  necessary  to  protect 
the  wires  in  the  test  tube ;  generally  it  is  sufficient  to 
spread  them  apart  to  prevent  short  circuiting  the 
current.  From  the  mouth  of  the  test  tube  the  wires 
may  or  may  not  be  protected  as  is  found  necessary 
or  convenient.  The  cold  junction  may  be  located  in 
a  bottle  with  a  thermometer  inserted  in  the  cork. 


134  TESTING  FOR  PROCESSES 

In  ordinary  work  this  is  a  useless  refinement  and 
the  temperature  of  the  cold  junction  may  be  taken 
as  that  of  the  surrounding  room.  Care  should  be  taken 
to  have  the  galvanometer  sufficiently  removed  from 
the  source  of  the  heat.  When  reading  the  deflec- 
tions for  the  boiling  points,  the  needle  of  the  gal- 
vanometer will  reach  a  stationary  position  when  the 


Pig.  20.     APPARATUS  FOR  THE  DETERMINATION  OF 
MELTING  AND  FREEZING  POINTS  OF  METALS 


thermo-couple  junction  has  reached  an  equilibrium 
of  temperatures  with  the  vapors.  The  pyrometer 
junction  should  be  placed  well  within  the  plaster 
muff  but  not  below  the  surface  of  the  liquid.  It 
is  well  to  let  the  liquid  cool  and  re-heat  several  times 
so  as  to  furnish  several  observations.  When  couples 
other  than  the  platinum  groups  are  standardized,  it 
is  well  to  protect  them  from  the  vapors  of  sulphur 


PYROMETRY  135 

by  means  of  a  sealed  hard  glass  tube  as  shown  in 
Fig.  19.  For  the  determination  of  the  melting  points 
or  freezing  points  of  metals  the  wires  should  always 
be  protected ;  the  best  covering  is  afforded  by  porce- 
lain tubes  as  shown  in  Fig.  20.  About  35  to  40  grams 
of  the  substance  to  be  used  is  taken  and  placed  in  a 
small  crucible.  No.  00  graphite  crucible  is  well 
suited  for  the  work.  For  zinc,  aluminum,  silver, 
and  copper  the  crucible  method  is  to  be  preferred, 
while  for  precious  metals,  gold  and  platinum,  the 
wire  method  is  the  better.  In  the  crucible  method 
an  apparatus  is  prepared  as  is  shown  in  Fig.  20  by 
taking  two  crucibles  fitting  snugly  one  within  the 
other.  Two  tuyere  holes  are  cut  in  the  outside  cru- 
cible and  the  blast  is  supplied  through  the  openings. 
The  apparatus  answers  also  well  for  the  melting  of 
gold  and  copper. 

When  making  determinations  it  is  well  to  insert 
the  porcelain  tubes  at  the  beginning  of  the  opera- 
tion, before  the  flame  is  applied,  so  as  to  avoid 
cracking  them  by  too  sudden  heating.  The  heating 
should  be  conducted  slowly  and  continued  until  the 
metal  is  melted;  this  can  be  ascertained  either  by 
direct  observation,  by  feeling  the  solid  metal  with 
the  porcelain  tubes,  or  by  the  deflection  of  the  gal- 
vanometer. It  is  not  necessary  in  the  crucible  meth- 
od to  watch  the  melting  point.  Once  certain  the 
metal  is  fused,  the  heating  is  stopped,  the  furnace 
is  allowed  to  cool  slowly  and  regularly  and  at  the 
instant  the  mass  has  reached  the  freezing  tempera- 
ture, a  well  marked  stop  of  the  galvanometer  needle 
is  observed.  With  proper  care  and  gradual  cooling 


136  TESTING  FOR  PROCESSES 

it  is  possible  to  have  the  needle  stationary  for  30 
seconds  with  a  mass  of  35  grams  of  metal.  It  is  al- 
ways well  to  repeat  the  operation. 

Temperature  According  to  Color. — Approximate 
temperatures  may  be  determined  by  observing  the 
color.  For  this  purpose  the  following  table  may  be 
used: 

Color.  Degrees  C. 

Lowest  red  visible  in  the  dark 470 

Dark  blood-red  or  black-red 532 

Dark  red,  blood-red,  low  red 566 

Dark  cherry-red  635 

Cherry-red,  full  red   746 

Light  cherry,  light  red 843 

Orange    900 

Yellow    1,000 

Light  yellow 1,080 

White    1,205 


RETORTING 

A  good  way  to  prepare  amalgam  for  retorting  is 
to  wash  and  grind  it  in  a  Wedgewood  mortar  to  re- 
move the  sand,  then  treat  it  with  dilute  nitric  acid 
to  remove  dirt  and  copper,  if  present.  A  few  pieces 
of  metallic  zinc  agitated  with  the  mercury  and  amal- 
gam will  often  prove  beneficial.  Strain  through  tick- 
ing or  chamois  skin  to  separate  the  solid  amalgam. 
The  retorts  are  of  various  shapes  and  sizes.  For  lab- 
oratory use  a  half  pint  size  as  shown  in  the  illustra- 
tion is  frequently  used.  Such  a  retort  consists  of  an 
iron  pot  (a)  with  a  cover  (b)  held  firmly  in  place 
by  a  clamp  (c)  and  a  set  screw  (d).  A  bent  iron  tube 
(e)  leads  off  the  fumes,  which  are  condensed  with 
the  aid  of  a  water  jacket  of  rags  (f)  and  collected 
under  water  in  a  dish. 

The  retort  should  be  thoroughly  cleaned  and  coated 
on  the  inside  with  chalk,  ruddle  (Fe203),  graphite, 
or  better  than  any  of  those,  first  chalked  on  the  in- 
side and  then  lined  with  paper.  This  will  prevent 
the  residue  from  sticking  to  the  bottom  of  the  re- 
tort in  case  the  heat  is  too  high  or  zinc  or  lead  is 
present.  Place  the  amalgam  in  the  retort,  not  over 
two-thirds  full,  smear  the  rim  with  a  thick  paste  of 
flour,  to  make  a  tight  joint,  and  clamp  on  the  cover 
tightly.  Heat  very  gently  at  first  and  gradually  in- 
crease the  temperature  but  never  above  a  dull  red, 
otherwise  the  retort  will  be  bulged  or  melted.  Mer- 
cury boils  at  674°F.  or  357°C.  No  fluxing  material 
like  borax  should  be  used  in  the  retort  or  the  spher- 


RETORTING  139 

oidal  mercury  will  be  changed  to  cohesive  and  boil 
with  great  violence.  The  residue  might  also  be  found 
permanently  braized  to  the  interior  of  the  retort. 
Do  not  dip  the  delivery  tube  deep  in  the  water.  Just 
immerse  the  end  of  it  in  a  small  dish  in  order  to  pre- 
vent the  possibility  of  water  being  sucked  back  into 
the  retort  by  any  sudden  cooling  and  thereby  caus- 
ing a  violent  explosion  because  of  the  sudden 
generation  of  steam  in  the  heated  retort.  Treat  the 
resulting  residue  as  given  under  'Refining.' 

For  small  amounts  of  mercury,  where  the  gold 
alone  is  wanted,  put  the  mercury-amalgam  in  a 
parting  flask,  after  thoroughly  cleaning,  and  dissolve 
off  the  mercury  with  nitric  acid  aided  by  heat.  When 
completely  parted,  wash  off  the  salts  of  mercury  and 
acid  with  distilled  water.  If  the  residue  contains 
foreign  matter,  filter  carefully,  so  as  not  to  lose  any 
of  the  gold  or  residue,  burn  the  filter  paper  in  a 
lead  tray  weighing  10  grams,  wrap  the  residue  in 
the  tray  with  enough  silver  to  inquart  the  gold  and 
cupel.  The  gold  is  now  recovered  by  parting.  If 
pure,  the  gold  may  be  annealed  and  weighed  at  once 
without  filtering  or  cupeling. 


REFINING  OF  BULLION 

If  the  bullion  is  base,  nitre  and  borax  glass  are  both 
needed  in  refining,  but  too  much  nitre  will  rapidly 
eat  into  the  graphite  crucible. 

Lead,  when  present  in  the  bullion,  is  best  oxidized 
by  nitre  or  sal-ammoniac;  tin,  by  means  of  potas- 
sium carbonate;  antimony  and  arsenic,  by  means  of 
nitre  or  by  stirring  the  bullion  with  an  iron  rod.  So- 
dium carbonate  is  to  be  avoided  unless  there  is  sili- 
cious  matter  present,  still  a  little  of  it  with  nitre 
seems  to  work  well  even  if  the  bullion  is  quite  pure. 
Bone-ash  and  silica  save  the  crucible  from  the  action 
of  the  oxides  and  are  especially  useful  in  thickening 
the  slags  in  case  thickening  is  necessary.  If  bullion 
is  poured  together  with  the  slag  it  should  be  per- 
fectly liquid. 

Toughening  serves  to  eliminate  small  quantities 
like  arsenic,  antimony,  and  lead,  which  render  the 
bullion  brittle  and  unfit  for  coinage  or  similar  pur- 
poses. T.  K.  Rose  recommends  the  addition  of  a  lit- 
tle sal-ammoniac  or  corrosive  sublimate  to  the  melted 
bullion.  Cover  quickly  to  keep  the  volatile  chloride 
fumes  out  of  the  room.  Test  by  casting  a  small  ingot, 
cooling  and  seeing  if  it  will  stand  bending  back  upon 
itself  without  cracking. 

Pure  gold  is  a  brilliant  green  color  when  melted 
and  it  may  then  be  poured.  Silver  when  nearly  pure 
often  bubbles  violently  in  the  crucible  and  especially 
so  when  much  nitre  has  been  used  in  the  refining. 
The  remedy  is  to  lower  the  temperature,  cover  the 


REFINING  141 

melt  with  charcoal  and  stir  with  a  graphite  rod  un- 
til bubbling  ceases,  then  pour. 

In  pouring  and  casting,  stir  thoroughly  and  take 
a  sample  for  assay  immediately  after.  Cast  in  heated 
molds  well  coated  with  oil.  Some  use  one-fourth  inch 
of  oil  in  the  bottom  of  the  mold,  others  sprinkle  rosin 
in  the  bottom  just  before  pouring.  The  object  of  the 
oil  is  to  make  a  smoother  ingot  and  by  its  burning 
on  the  top  of  the  ingot  to  prevent  all  sprouting  and 
tarnishing.  Pouring  small  quantities  of  bullion  is 
generally  unsatisfactory  and  better  results  may  be 
attained  by  letting  the  crucible  and  content  cool. 
Break  the  crucible  and  clean  the  button,  saving  the 
slag  for  further  treatment.  Very  small  amounts  of 
bullion  may  be  refined  by  fusing  with  a  charge  as 
used  in  the  regular  assay  and  cupelling  the  resulting 
lead  button. 

In  treating  cyanide  clean-ups,  two  methods  are 
available. 

1.  Mix  the  wet  slime  with  a  little  nitre,  place  on 
a  tray  in  a  muffle  furnace  and  when  dry  increase  the 
heat  sufficiently  to  oxidize  a  large  portion  of  the  zinc. 
The  oxidized  charge  may  be  smelted  with  the  follow- 
ing flux : 

Parts. 

Dry  slime   100 

Bicarbonate  of  soda 40 

Borax 40 

Silica    15 

The  bullion  smelted  by  this  method  may  be  800 
fine. 

2.  In  the  acid  treatment  the  zinc  is  dissolved  in 


142  TESTING  FOR  PROCESSES 

a  vat  by  adding  sulphuric  acid  until  all  action  has 
ceased  when  aided  by  heat.  The  slime  is  washed 
by  agitating  with  boiling  water  and  collecting  in  a 
filter-press.  The  following  flux  is  an  example  of  one 
used  in  smelting  air-dried  slime : 

Parts. 

Slime    100 

Borax    50 

Bicarbonate   of   soda 30 

The  following  is  an  example  of  a  flux  used  when 
the  slime  contained  lead  from  the  use  of  the  zinc- 
lead  couple : 

Per  cent. 

Borax    60.0 

Nitre    19.0 

Sand    11.5 

Soda    7.0 

Manganese  dioxide  may  be  used  to  replace  nitre, 
as  it  has  a  higher  oxidizing  power  than  the  latter. 
Its  use  is  not  to  be  recommended  with  highly  argen- 
tiferous slime  as  it  tends  to  enrich  the  slag  in  silver. 

Other  methods  used  are  scorification-  and  cupella- 
tion  in  an  English  cupelling  furnace;  smelting  in  a 
small  blast  furnace  to  base  bullion;  retorting  the 
zinc  and  fusion  of  the  retort  residues. 


CONCENTRATION  TESTS 

It  will  be  seen  in  the  preceding  chapters  that  con- 
centration will  often  be  a  necessary  part  of  an  ore- 
treatment  scheme.  The  following  suggestions  will 
prove  of  value  in  determining  the  methods  to  be  fol- 
lowed. 

PRELIMINARY  OBSERVATIONS 

Value  and  Character  of  Constituent  Mineral. — 
Take  a  50-gram  sample  of  the  ore  and  crush  until  the 
concentrate  is  free  from  the  gangue.  Separate  the 
minerals  from  the  gangue  by  hand-picking,  panning, 
by  using  magnets,  or  with  heavy  solutions  as  the 
case  may  require.  Each  mineral  taken  out  should  be 
entirely  free  from  gangue  or  other  minerals.  The 
total  amount  of  each  constituent  is  determined  by 
analysis  but  the  approximate  amounts  may  be  found 
from  the  weights  separated.  Assay  each  mineral  for 
gold,  silver,  and  constituent  metal.  It  will  prove 
useful  to  determine  the  specific  gravities  as  well  as 
to  note  the  shapes  and  other  characteristics  of  the 
grains.  The  value  of  the  mineral  to  be  saved  will 
limit  the  closeness  of  the  concentration  and  the  elab- 
orateness of  the  scheme  to  be  followed.  If  the  dif- 
ference in  specific  gravity  of  the  minerals  to  be  sepa- 
rated is  as  much  as  two  points,  no  especial  difficulty 
will  be  found  in  the  usual  methods  of  water  concen- 
tration, other  conditions  being  favorable.  If  the 
products  to  be  separated  are  of  the  same  specific  gra- 
vity, other  methods  must  be  followed.  Zinc  blende 


144  TESTING  FOR  PROCESSES 

may  be  separated  from  garnet,  rhodonite,  and  sider- 
ite  by  the  Elmore  vacuum  oil  process.  Pyrite  is  re- 
movable from  zinc  blende  by  giving  the  mixture  a 
flash  roast  and  taking  out  the  iron  with  a  magnetic 
separator.  In  general,  the  closer  the  two  minerals 
to  be  separated  are  in  specific  gravity,  the  closer  must 
be  the  sizing  for  jigs  and  the  classification  for  tables, 
and  the  more  attention  must  be  paid  to  re-crushing 
and  re-treating  the  middling. 

Friable  minerals  in  general  demand  graded  crush- 
ing, and  graded  concentrating  or  the  removal  of  con- 
centrate as  soon  as  a  sufficient  quantity  is  liberated, 
as  little  screening  as  is  consistent  with  good  work, 
and  the  use  of  free-crushing  machinery  such  as  rolls. 
In  many  cases  the  valuable  mineral  is  the  most  friable 
constituent  of  the  ore,  hence  the  necessity  of  free- 
crushing  conditions.  By  graded  crushing  is  meant 
the  reduction  of  the  lump  ore  to  the  final  sizes  by 
several  stages  of  crushing  and  the  removal  of  all  un- 
dersize  from  the  feed  of  each  crusher.  Screening 
can,  however,  be  carried  too  far  as  sizing  by  this 
method  produces  slime  by  abrasion.  By  free-crush- 
ing conditions  is  meant  the  use  of  rolls,  on  account 
of  the  small  amount  of  slime  that  they  produce  in 
comparison  with  other  crushers.  In  other  crushing 
devices,  such  as  stamps,  the  conditions  may  be  made 
very  favorable  by  using  a  large  screen  discharge 
area  with  abundant  water  and  low  height  of  dis- 
charge. Concentrate  should  be  removed  as  soon  as 
liberated  as  further  crushing  will  unnecessarily  slime 
it,  also  because  coarse  concentration  is  cheaper  and 
more  effective  than  fine. 


CONCENTRATION  TESTS  145 

Determination  of  Included  Grain. — Crush  seven 
samples  of  the  ore  so  as  to  pass  through  screens  of 
the  following  sizes :  4,  8,  16,  20,  40,  80,  and  100  mesh. 
Minerals  may  be  liberated  in  sizes  coarser  than  4- 
mesh  and,  if  this  seems  probable,  larger  sizes  should 
be  tried.  Separate  the  concentrate  by  hand-picking 
or  panning  until  entirely  free  from  gangue  in  each 
case. 

With  the  percentages  of  concentrate  obtained  in 
the  above  tests  as  ordinates,  and  the  corresponding 
meshes  as  abscissae,  plot  a  curve  which  will  clearly 
show  at  what  mesh  the  maximum  amount  of  concen- 
trate is  liberated.  The  curve  will  show  at  what  stage 
of  the  crushing  to  introduce  concentration,  how  much 
finer  the  ore  must  be  re-crushed  before  introducing 
another  concentrating  unit,  to  what  degree  of  fineness 
all  tailing  or  middling  must  be  crushed  to  completely 
liberate  the  included  grains,  the  probable  recovery, 
and  the  percentage  of  concentrate  in  the  ore.  Addi- 
tional value  would  be  given  to  the  above  graph  if 
the  crushed  samples  were  separated  into  clean  con- 
centrate, middling  and  clean  tailing,  and  the  data 
incorporated  in  the  curve.  A  microscope  will  prove 
valuable  in  determining  the  percentages  in  the  finer 
sizes. 

SCKEEN  SIZING  TESTS 

In  the  examination  of  ores  and  mill  products,  it 
is  often  necessary  to  know  the  percentages  of  the 
different  sizes  present  as  well  as  their  assay  values. 
This  may  be  learned  by  making  a  screen  test.  The 
screening  is  done  by  placing  a  weighed  sample  of  the 


146  TESTING  FOR  PROCESSES 

ore  or  product,  say  100  to  500  grams,  in  the  upper 
screen  of  a  nest  arranged  in  order  of  size,  with  the 
coarsest  on  top,  and  giving  the  whole  bank  a  jarring, 
rocking  motion.  Before  taking  any  screen  from 
the  nest  it  should  be  shaken  separately  for  a  definite 
length  of  time  to  make  the  last  of  the  undersized 
grains  pass  through.  In  case  of  the  finer  sizes  to 
start  with,  only  a  small  portion  of  the  ore  should  be 
placed  on  the  screen  at  one  time,  because  overcrowd- 
ing destroys  the  efficiency  of  the  screening.  The  ore 
resting  on  each  sieve  is  weighed  and  if  values  are 
wanted,  it  is  ground,  if  less  than  100  mesh,  mixed, 
sampled  and  assayed.  The  sizes  of  screens  used 
might  conveniently  be  4,  8,  16,  20,  30,  40,  60,  80,  100, 
150,  and  200  mesh.  These  tests  will  be  of  value  in 
finding  where,  or  in  what  size  the  value  lies  in  tailing 
or  other  mill  products  and  thus  suggesting  a  method 
of  obtaining  a  higher  extraction.  They  may  be  of 
use  in  determining  the  proportion  of  crushing  to  con- 
centrating machinery,  also  in  determining  the  effici- 
ency of  crushers,  grinders,  and  screens.  In  consid- 
ering the  re-treatment  of  table  middling  resulting 
from  a  classified  feed,  a  screen  test  will  show  if  it 
would  be  of  advantage  to  screen  out  and  reject  the 
larger  and  necessarily  low-grade  particles  before  fur- 
ther concentration  or  treatment. 

WET  CONCENTRATION 

Preliminary  Tests. — Concentrate  coarser  than 
94-in.  in  diameter  may  be  efficiently  removed  by 
hand-picking.  Waste  rock  may  be  removed  in  the 
same  operation.  For  sizes  down  to  12  or  16  mesh 


CONCENTRATION  TESTS  147 

a  jig  will  be  the  most  efficient  means  of  recovering 
the  concentrate  but  on  the  finer  sizes  a  table  does 
better  work. 

Jigging  Tests. — To  feed  a  jig,  so  as  to  obtain  the 
most  efficient  work  the  material  should  be  screen- 
sized  or  at  least  hydraulically  classified.  On  account 
of  the  metallic  minerals  being  present  in  the  classi- 
fied feed  in  smaller  particles  than  those  of  the  gangue, 
the  product  is  better  suited  to  hutch  work  when  jig- 
ging. The  proper  range  of  sizes  to  be  treated  on  one 
jig  is  limited  by  the  hindered  settling  ratios  of  the 
minerals  to  be  separated. 

HINDEBED  SETTLING  RATIOS  OF  QUARTZ  AND  COMMON  MINERALS* 

Quartz  Ratio  diam.  quartz, 

and —  to  diam.  mineral. 

Copper 8.6 

Galena    5.8 

Wolframite  5.2 

Antimony    4.9 

Cassiterite   4.7 

Arsenopyrite   3.7 

Chalcocite    3.1 

Pyrrohotite    2.8 

Sphalerite 2.1 

Epidote  2.0 

Anthracite   0.2 

*Taken  from  Richards'  'Ore  Dressing/  Vol.  I. 

Take  as  an  example  the  minerals  quartz  and  ga- 
lena of  which  the  hindered  settling  ratio  is  5.8.  This 
means  that  a  particle  of  quartz  5.8  times  as  large  as 
a  particle  of  galena  will  be  found  in  the  same  posi- 
tion or  layer  in  a  jig  bed.  Hence,  to  obtain  a  perfect 
separation  of  clean  concentrate  from  clean  tailing, 


148  TESTING  FOR  PROCESSES 

the  range  of  sizes  to  be  treated  on  one  jig  must  be 
somewhat  less  than  the  hindered  settling  ratio.  Mid- 
dling or  included  grains  will  assume  positions  accord- 
ing to  the  percentage  of  included  concentrate  and 
must  be  re-crushed  and  re-sized  for  further  treat- 
ment. In  the  coarser  sizes,  all  the  tailing  and  mid- 
dling may  require  re-crushing,  as  is  the  case  with 
some  of  the  Butte  ores. 

In  running  a  jig  for  concentrate  only,  a  screen 
bed  is  used  in  which  the  opening  is  slightly  smaller 
than  the  smallest  grain  in  the  feed.  With  the  finer 
sizes,  10  to  20  mesh,  a  16-mesh  screen  would  be  used 
with  a  i/^-in.  bed  of  coarse  concentrate  (4  to  10  mesh), 
for  the  purpose  of  making  hutch  product  and  con- 
centrate. In  separating  the  finer  sized  particles, 
screening  will  often  prove  impractical,  in  which  case, 
a  hydraulic  classifier,  acting  under  free  settling  con- 
ditions, should  replace  the  last  screen  or  trommel. 

A  25  to  50-lb.  sample  of  the  ore  is  crushed  as 
nearly  as  possible  under  conditions  that  would  be 
used  in  practice.  The  coarser  sizes  are  screened  out 
for  the  jig  work  and  the  undersize  saved  for  further 
treatment.  Each  lot  or  size,  which  has  been  pre- 
viously weighed  and  assayed,  is  run  over  a  jig  of 
the  Vezin  laboratory  size  and  the  material  separated 
into  concentrate,  middling,  hutch  product  (if  any), 
and  tailing.  Weigh  and  assay  each.  The  assay 
values  of  the  middling  and  tailing  will  determine 
whether  they  are  to  be  rejected  or  treated  further. 

Fine  Concentration. — This  work  may  be  done 
roughly  in  a  miner's  gold  pan  or  under  conditions 
more  nearly  resembling  those  in  a  mill  on  a  labora- 


CONCENTRATION  TESTS  149 

tory  size  table  of  the  Wilfley  type.  The  products 
treated  may  include  the  undersize  from  the  trommels 
whose  oversize  is  fed  to  jigs,  re-ground  middling  or 
tailing  from  the  fine  jigs,  re-ground  table  middling, 
or  all  the  product,  if  it  has  been  crushed  at  once  finer 
than  16  mesh.  To  secure  the  best  results,  the  feed 
to  a  table  should  be  hydraulically  classified  into  two 
or  more  sizes  according  to  the  coarseness  of  the  ori- 
ginal feed  or  the  difficulty  of  the  separation.  As 
many  as  six  sizes  are  used.  The  pulp  should  not  be 
too  liquid,  hence  it  should  be  de-watered  before  send- 
ing to  the  classifiers  or  feeding  to  the  tables.  The 
overflow  from  the  classifiers  and  the  de-watering 
boxes  should  go  to  a  final  large  settling  tank  whose 
spigot  product  is  treated  on  slime  tables.  Each  lot 
or  size,  which  has  been  previously  weighed  and  as- 
sayed, is  run  over  a  laboratory  table  of  the  Wilfley 
type  and  separated  into  concentrate,  middling,  and 
tailing.  Weigh  and  assay  each.  The  further  treat- 
ment of  the  middling  is  dependent  upon  its  contained 
values. 

The  tailing  from  the  tables  treating  the  finer  sizes 
as  well  as  the  overflow  from  the  settling  tank,  may 
be  further  treated  by  running  over  a  small  canvas 
table.  One  10  ft.  long  and  2  to  4  ft.  wide,  covered 
with  No.  6  canvas  duck,  will  be  found  convenient. 
The  woof  or  cross  threads  are  laid  down  the  slope 
as  the  concentrate  is  more  effectually  caught  in  the 
higher  ridges  by  the  warp. 

Draw  up  a  report  which  will  include  the  per- 
centage of  recovery  of  each  separate  run  or  test 
based  on,  (a)  the  total  value  of  the  feed  to  each 


150 


TESTING  FOR  PROCESSES 


test,  (b)  the  total  value  of  the  total  weight  of  ore 
taken  for  the  experiment.  The  total  values  ac- 
counted for  in  the  total  concentrate  and  the  tailing 
subtracted  from  100%  will  give  the  losses  due  to 
slime  and  handling.  Give  the  probable  final  recovery 
and  a  flow  sheet  of  the  treatment  followed. 

MAGNETIC   SEPARATION 

This  process  is  based  on  the  fact  that  certain  min- 
erals possess  magnetic  qualities  which  permit  their 
being  lifted  out  of  the  attendant  gangue  by  the  ac- 
tion of  electromagnets.  By  roasting  at  a  dull  red 


Fig.  22.     ELECTROMAGNET  FOR  TESTING  ORE 

(After  Lr.   H.  L.   Huddart,    The  Engineering  and  Mining 
Journal) 

heat  for  a  short  time,  certain  non-magnetic  iron  min- 
erals may  be  rendered  magnetic  and  thus  a  separa- 
tion may  be  effected.  This  fact  is  used  in  the  sepa- 
ration of  pyrite  from  zinc  blende,  which  are  taken 
off  together  from  the  concentrating  tables  and  jigs. 


CONCENTRATION  TESTS  161 

For  effective  work  the  ore  fed  to  the  separator  must 
be  sized.  The  closeness  of  sizing  depends  upon  the 
similarity  of  the  magnetic  properties  of  the  minerals 
to  be  separated. 

Test. — Assay  the  original  ore  for  gold,  silver  and 
metal  constituents.  Take  a  definite  weight  of  sam- 
ple, say  500  grams,  place  in  a  cast-iron  roasting-dish 
and  heat  in  a  muffle  or  oven-furnace  at  a  dull  red 
heat  until  the  flames  of  burning  sulphur  have  entirely 
disappeared.  Cool  down  by  withdrawing  from  the 
furnace,  and  weigh.  Note  the  change  in  weight  due 
to  the  roast.*  Remove  the  magnetic  mineral  by 
carefully  passing  over  the  machine  shown  in  Fig.  22. 
The  material  should  be  treated  two  or  three  times. 
The  separation  may  be  effected  by  spreading  out 
the  roasted  ore  on  a  clean  level  sheet  of  paper  in  a 
thin  layer  and  removing  the  magnetic  mineral  with 
a  strong  horseshoe  magnet.  Weigh  and  assay  both 
the  concentrate  and  the  tailing.  Calculate  the  per- 
centage of  recovery  or  the  effectiveness  of  separa- 
tion as  the  case  may  require. 

ELECTROSTATIC  SEPARATION 

This  process  is  based  on  the  difference  in  electrical 
conductivity  and  inductivity  of  minerals  in  an  elec- 
trical field,  and  also  the  fact  that  bodies  charged 
with  like  signs,  or  kind  of  electricity,  repel,  and  with 
unlike  signs  attract  each  other.  Every  mineral  if 
subjected  to  a  sufficiently  high  electrical  pressure  be- 
comes a  conductor  to  an  extent  depending  upon  the 

"This  roast  may  be  omitted  if  the  constituent  to  be  sep- 
arated is  naturally  magnetic. 


152  TESTING  FOR  PROCESSES 

mineral.  The  fundamental  principles  may  be  utilized 
in  various  ways,  of  which  two  are  given:  (A)  The 
mineral  mixture,  in  a  neutral  electrical  condition,  is 
brought  into  contact  with  a  highly  charged  surface, 
whereupon  the  best  conductors  are  expelled  first  and 
poorest  last.  (B)  The  mineral  mixture  with  a  charge 
of  one  kind  of  electricity  is  brought  into  contact 
with  a  surface  having  a  charge  of  the  opposite  sign, 
whereupon  the  good  conductors  will  leave  the  sur- 
face first,  and  the  poorest  is  attracted  and  held  the 
longest,  by  reason  of  its  reluctance  to  change  its 
charge  of  an  unlike  sign  to  a  like  sign. 

The  following  is  a  list  of  minerals  placed  accord- 
ing to  their  conductivities.  This  list  indicates  clearly 
the  possibilities  of  the  process,  the  basic  principle 
stated  first  being  recalled. 

GOOD  CONDUCTOBS  POOR  CONDUCTORS 

Most  sulphides  Quartz 

Pyrite  Quartzite 

Chalcopyrite  Sandstone 

Galena  Feldspars 

Native  metals  Granite 

Copper  Porphyry 

Gold  Andesite 

Some  oxides  Epidote 

Magnetite  Garnet 

Hematite  Calamine 

Most  carbonates 
Calcite 
Siderite 
Limestone 
Most  sulphates 
Barite 
Gypsum 
Sphalerite  (zinc  blende) 


CONCENTRATION  TESTS  153 

The  principal  application  of  this  process  is,  at 
present,  the  separation  of  zinc  blende  from  other 
sulphides  where  water  concentration  fails  on  account 
of  the  slight  differences  in  specific  gravity.  There 
may  be  thus  produced  from  a  zinciferous  concen- 
trate of  low  or  doubtful  value  because  of  the  pres- 
ence of  zinc,  a  high-grade  zinc  product  and  a  very 
much  improved  concentrate  of  the  remaining  sul- 
phides. Electrostatic  separation  might  also  be  used 
to  effect  a  dry  concentration  of  sulphide  ores  from 
quartz  and  other  gangues,  where  water  is  scarce; 
for  the  concentration  of  easily  slimed  ores,  such  as 
the  silver  sulphides;  for  the  separation  of  high 
specific  gravity  minerals,  such  as  garnet,  barite,  and 
epidote,  from  the  valuable  sulphides ;  for  the  concen- 
tration of  molybdenite,  graphite,  and  monazite,  as 
well  as  other  minerals  where  the  differences  in  elec- 
trical conductivities  open  a  way. 

The  process  should  be  taken  into  account  where 
the  simpler  methods  of  water  concentration  are  not 
successful  because  of  commercial  reasons,  such  as 
water  supply  or  because  of  the  physical  nature  of 
the  mineral.  It  may  also  be  used  in  conjunction 
with  water  concentration. 

The  experiments  should  be  conducted  with  a  com- 
mercial sized  electrostatic  unit,  under  such  rules  as 
would  govern  a  regular  concentration  test.  The 
Huff  Electrostatic  Separator  Co.  make  a  laboratory 
unit  for  testing  3-pound  lots  of  ore. 

OIL  FLOTATION 

The  Elmore  vacuum  process  is  based  primarily 


154  TESTING  FOR  PROCESSES 

upon  the  fact  that  oil  has  the  power  of  wetting  cer- 
tain metallic  mineral  particles  as  distinct  from  the 
gangue.  The  selective  action  is  further  increased 
in  some  cases  by  the  fact  that  gas  bubbles  liberated 
by  an  acid  and  aided  by  a  partial  vacuum,  attach 
themselves  to  the  greased  particles  thus  causing  them 
to  float.  The  process  is  applicable  to  the  separation 
of  chalcopyrite  from  magnetite  and  spathic  iron 
gangue,  galena  and  zinc  blende  from  baryta,  garnet 
and  similar  heavy  gangues.  Friable  minerals  such 
as  sulphides  of  antimony  and  molybdenum  that  suf- 
fer high  losses  in  water  concentration  are  found 
amenable  to  concentration  by  this  process.  The 
process  does  not  lend  itself  readily  to  simple  and  in- 
expensive laboratory  experiments.  Samples  should 
be  shipped  to  the  manufacturers  of  the  apparatus 
for  testing. 


DETERMINATION  OF  THE  SUITABILITY  OF  AN 
ORE  FOR  SMELTING 

Preliminary  estimates  of  the  profit  or  loss  in 
smelting  an  ore  are  conveniently  made  by  applying 
a  schedule  of  smelter  charges  that  would  be  used 
for  the  territory  in  which  the  ore  occurs.  This 
method  does  not  require  the  determination  of  the 
most  economical  blast-furnace  charge  for  smelting 
the  ore  under  the  conditions  existing  in  the  region  in 
question  or  the  cost  of  fluxes,  fuel,  labor  and  con- 
struction of  a  plant,  since  all  these  items  are  embod- 
ied in  the  smelter  schedule  as  the  results  of  expe- 
rience with  the  ores  in  the  vicinity.  In  cases  where 
there  is  no  competition,  the  charges  and  penalties 
might  be  excessively  high  and  the  payments  for 
metals  correspondingly  low,  which  must  be  care- 
fully considered  when  making  estimates.  In  figur- 
ing the  final  profit,  allowances  must  be  made  for 
mining  and  handling. 

The  schedule  may  be  used  for  determining  the 
advisability  of  smelting  the  ore  raw  or  concentrat- 
ing first,  as  well  as  the  limit  or  cleanness  -of  con- 
centration, by  applying  the  schedule  to  a  given 
amount  of  raw  ore,  say  100  tons,  and  to  the  concen- 
trate obtained  from  an  equal  amount  of  ore.  In  the 
same  way  the  limit  of  concentration  may  be  deter- 
mined. An  analysis  of  a  fair  sample  must  be 
made  for  gold,  silver  and  other  substances  that  are 
penalized  or  rewarded  under  the  proper  schedule. 


156  TESTING  FOR  PROCESSES 

SCHEDULE 

FOB    DRY    ORES,     CONCENTRATE    AND    'TAILING/    LEAD    ORES,    AND 

LEAD  CONCENTRATE,   CLEAR  CREEK  AND   GILPIN   COUNTIES, 

COLORADO 

February  1,  1905 

All  rates  f.o.b.  cars,  Denver. 

DRY  'TAILING'  AND  CONCENTRATE 

GOLD,  $19  per  oz.,  if  5/100  oz.  or  over  per  ton. 

SILVER,  95%  of  N.  Y.  quotation  day  of  assay,  if  1  oz.  or 

over  per  ton. 
COPPER,  dry  assay  (wet  less  1.5  units).*  Per  unit. 

5%  or  less $1.25 

Over  5  and  including  10% 1.50 

Over  10% 1.75 

10%  silica  basis,  lOc.  charge  for  each  per  cent  over  lO.f 

5%  zinc  basis,  30c.  charge  for  each  per  cent  over  5. 

When  gross  value  is:  Treatment  charge. 

Not  over  $35  per  ton $3.50 

Over  $35  and  including  $80  per  ton 4.00 

Over  $80  per  ton 5.00 

Upon  lots  of  less  than  7  tons 5.00 

*A  unit  is  1  per  cent. 

fSubtract  silica  from  iron,  i.  e.,  an  ore  containing  30% 
SiO2  and  15%  Fe  would  be  figured  at  30  —  15  =  15%  SiO2, 
and  vice  versa. 


DRY  SILICIOUS  AND  COPPER  ORES 

GOLD,  $19.50  per  oz.,  if  5/100  oz.  or  over  per  ton. 

SILVER,  95%  of  N.  Y.  quotation,  day  of  assay. 

COPPER  as  in  schedule  of  concentrate. 

$8  treatment  charges. 

40%  silica  basis,  5c.  credit  for  each  per  cent  less  than  40, 
and  lOc.  charge  for  each  per  cent  over  40,  up  to  a  maxi- 
mum charge  of  $11  on  ores  not  exceeding  $25  gross 
value;  and  $12.50  on  ore  exceeding  $25  gross  value. 

5%  zinc  limit,  30c.  charge  for  each  per  cent  over  5. 


SMELTING  ORES  157 

OXIDIZED    IRONY    ORES 

GOLD,  $19  per  oz.,  if  5/100  oz.  or  over  per  ton. 
SILVER,  95%  of  N.  Y.  quotation,  date  of  assay. 
LEAD,  25c.  credit  per  unit  for  5%  or  over. 
$2  treatment  charges. 

Neutral  basis,  lOc.  per  unit  charge  for  each  per  cent  silica 
excess. 

LEAD   ORES 

GOLD,  $19.50  per  oz.,  if  B/100  oz.  or  over  per  ton. 

SILVER,  95%  of  N.  Y.  quotation  date  of  assay. 

LEAD,  prices  flat. 

COPPER,  $1  per  unit  dry  (1.5%  off  wet)   when  ore  assays 

2%  wet. 
ZINC  limit  10%,  50c.  charge  for  each  per  cent  over  10. 


NEUTRAL   SCHEDULE 


Assay,  p 
Pb  incl 
5 
Over  10 

"  15 
"  20 
"  25 
"  30 
"  35 
"  40 
"  45 
"  50 

er  cent 
usive. 
to  10 

Cents 
per  unit. 
25 

W.C.* 

$8.00 
7.00 
5.00 
4.00 
4.00 
3.00 
2.50 
2.00 
2.00 
2.00 

"  15 

25 

"  20   

25 

"  25  

25 

"  30 

30 

"  35 

30 

"  40  

30 

"  45  

32 

"  50 

35 

40 

.  Neutral  basis,  lOc.  credit  for  each  per  cent  iron  excess, 
and  lOc.  charge  for  each  per  cent  silica  excess. 

*Working  charge. 


158 


TESTING  FOR  PROCESSES 


FLAT   SCHEDULE 


Assay,  p 
Pb  incl 
5 
Over  10 

"  15 
"  20 
"  25 
"  30 
"  35 
"  40 
"  45 
"  50 

er  cent 
usive. 
to  10  

Cents 
per  unit. 
.  .  .  .       25 

w.c. 

$12.00 
10.50 
8.50 
6.50 
6.00 
4.50 
3.00 
2.00 
2.00 
2.00 

"  15  

25 

"  20 

25 

"25  

25 

"  30  

30 

"  35 

30 

"  40  

.   ...       30 

"  45  

32 

"  50 

35 

40 

Neutral  schedule  to  be  used  when  it  figures  better  for  the 
shipper. 


LEAD   CONCENTBATE 

GOLD,  $19  per  oz.,  if  5/100  oz.  or  over  per  ton. 

SILVER  and  COPPER  as  in  lead  ores. 

LEAD,  prices  flat. 

SILICA  limit,  10%,  lOc.  charge  for  each  per  cent  silica 

excess  over  10. 
ZINC  limit,  5%,  30c.  charge  for  each  per  cent  zinc  excess 

over  5. 


Assay,  per  cent 
Pb  inclusive. 


Cents 
per  unit. 


5  to  10  25 

Over  10   "    15  24 

"  -  15    "    20  30 

"     20    "    25  32 

"     25    "    30  .  35 


W.C. 

$4.75 
4.00 
3.50 
3.25 
3.25 


Upon  concentrates  assaying  over  30%  lead,  apply  'Neutral 
schedule'  or  'Flat  schedule/  whichever  figures  better  for  the 
shipper;  $19  for  gold. 


SMELTING  ORES  159 


SCHEDULE 

FOB    ORES    AND    CONCENTRATE,    BOULDER    COUNTY     AND    CBIPPLE 
CREEK,   COLORADO 

February  1905 
All  rates  f.o.b.  cars,  Denver. 
GOLD,  $19  per  oz.,  B/wo  to  2  oz.  inclusive  per  ton. 
GOLD,  $19.50  per  oz.,  if  over  2  oz.  per  ton. 

SILVER,  90%  of  N.  Y.  quotation  if  ore  assays  from  1  to  10 

oz.  per  ton. 
SILVER,  95%  of  N.  Y.  quotation  if  ore  assays  over  10  oz. 

per  ton. 

Gross  value.  Treatment. 

Up  to  $10            $4.00 

Over      10  to  20 5.00 

20    "    30 5.50 

30    "    40 6.00 

40    "    50 6.50 

60    "    75 7.00 

75    "100 8.00 

"       100             9.00 

When  ore  does  not  exceed  $10  gross  value,  3%,  sulphur 
limit,  25c.  charge  for  each  per  cent  of  sulphur  over  3,  up  to 
a  maximum  charge  of  $2.50  per  ton,  zinc  limit  5%,  then  30c. 
charge  for  each  per  cent  of  zinc  over  5. 

When  ore  is  over  $10  gross  value,  no  sulphur  limit,  zinc 
limit  5%,  30c.  charge  for  each  per  cent  of  zinc  over  limit. 

LEAD   ORES 

Apply  schedule  for  lead  ores  for  Clear  Creek  and  Gilpin 
counties. 

CONCENTRATE,   LEAD   OR  DRY 

Apply  corresponding  schedule  for  Clear  Creek  and  Gilpin 
counties. 


PHYSICAL  PROPERTIES  OF  SLAGS 

General. — The  melting  point  of  a  slag  and  the 
superheat  necessary  above  the  melting  point  to  make 
it  flow  properly  may  be  determined  by  experiment. 
A  complete  analysis  of  the  slag  should  be  made  and 
the  silicate  degree  determined.  The  physical  charac- 
teristics visible  to  the  eye,  that  will  aid  in  the  ap- 
proximation of  the  analysis,  are  also  to  be  investi- 
gated. 

Take  a  500-gram  sample  of  the  granulated  slag 
and  make  a  complete  analysis.  Note  the  luster  and 
appearance  of  surface  and  fractures  of  the  granules. 
Determine  the  specific  gravity. 

Melting  Point  Determination. — Take  a  50-gram 
sample  and  grind  so  as  to  pass  a  100-mesh  sieve, 
being  careful  not  to  contaminate  with  foreign  sub- 
stances. Moisten  thoroughly  with  a  heavy  lubricat- 
ing oil  in  which  10  to  15%  vaseline  has  been  dis- 
solved by  the  aid  of  heat.  Ram  the  material  into  a 
slag  cone  mold  as  shown  in  Fig.  23,  in  thin  layers, 
scratching  each  surface  before  adding  the  next  lot  to 
provide  a  good  bond.  Use  a  rammer  %  in.  square 
and  ram  each  layer  very  hard  before  adding  the  next. 
Remove  the  cone  from  the  mould  by  unclamping  and 
sliding  the  halves  past  one  another.  Dry  by  setting 
in  front  of  a  muffle  and  raise  slowly  to  a  dull  red 
heat,  burning  off  the  binding  oil.  The  cone  is  now 
ready  for  testing.  Place  the  dried  cone  on  a  chrom- 
ite  or  graphite  slab  which  may  be  conveniently 


PROPERTIES  OF  SLAGS  161 

bedded  in  bone  ash  in  a  scorifying  dish  for  handling, 
and  place  in  the  interior  muffle  of  the  gas  forge. 
The  doorways  with  peep-holes  are  placed  in  position 
over  the  mouth  of  the  muffle  of  the  gas  forge.  The 
pyrometer  is  next  inserted  through  the  top  holes  so 
as  to  have  the  hot  junction  as  near  the  slag  cone  as 
possible. 

Light  the  gas  and  regulate  the  blast  so  as  to  take 
one-half  hour  in  bringing  the  muffle  to  a  dull  red 


Fig.  23.     CRUCIBLE  FOR  SLAG  TESTS 

a  Thin  fire-clay  tube;  6  thermocouple;   c  hot  junction; 

d  asbestos  plugs. 

heat.  In  the  next  hour  raise  slowly  past  1000 °C.  to 
the  melting  point  of  the  cones,  observing  the  melting 
temperature  of  the  cones  by  looking  through  the 
peep-holes.  Only  remove  the  inner  doors  for  obser- 
vation if  it  is  impossible  to  see  the  cones  on  account 
of  the  uniform  color  and  temperature  of  the  slag 
cones  and  interior  of  the  muffle. 


162 


TESTING  FOR  PROCESSES 


Necessary  Superheat  and  Appearance  When 
Molten. — Fill  with  granulated  slag  a  fire-clay  cruci- 
ble, lined  with  a  coating  of  chromite  and  water-glass ; 
if  necessary  on  account  of  a  basic  slag,  melt  down  in 
a  gas  forge-furnace,  until  thoroughly  liquid.  Dip  a 
cold  iron  rod  into  the  slag  and  note  the  appearance 
of  the  drops  falling  from  the  rod,  that  is,  whether 
they  string  out  or  break  off  short.  Note  the  appear- 
ance of  a  stream  of  slag  as  it  is  being  poured  into 
a  bucket  of  water.  By  means  of  several  trials  note 
the  temperature  at  which  the  molten  slag  is  liquid 
and  flows  properly.  The  temperature  of  the  slag  is 
taken  by  inserting  the  hot  junction  of  the  pyrometer, 
protected  by  a  thin  clay  tube,  well  into  the  slag  at 
the  beginning  of  the  experiment  and  having  an  ob- 
server read  the  temperature  indicated  by  the  gal- 
vanometer, whenever  so  directed  by  the  one  in 
charge  of  the  experiment.  The  temperatures  are  to 


Fig.  23.     SLAG  CONE  MOLD 

be  recorded,  with  appropriate  remarks.  Pour  some 
slag  into  a  mould  and  note  appearance  of  surface 
and  fracture. 

Report. — (1)  analysis  of  slag ;  (2)  specific  gravity; 
(3)    silicate   degree;    (4)    melting  temperature,   T\; 


PROPERTIES  OF  SLAGS  163 

(5)  temperature  of  slag  when  liquid,  T2 ;  (6)  super- 
heat; (7)  appearance  of  surface  of  molten  slag  when 
pouring,  also  nature  of  slag,  comparative  viscosity, 
tests  with  cold  iron  rod,  and  general  remarks;  (8) 
give  commercial  reasons  and  conditions  that  would 
require  the  use  of  this  slag. 

This  test  may  be  varied  to  show  the  characteristics 
for  any  particular  type  of  slag,  irony,  zincy,  limy, 
aluminous,  acid,  or  basic. 


SMELTING  PROPERTIES   OF   COPPER,   LEAD, 
GOLD,  AND  SILVER  ORES 

The  object  of  this  chapter  is  to  answer  the  ques- 
tion: Given  an  ore  or  collection  of  ores,  what  are 
the  metallurgical  conditions  to  be  attained  to  result 
in  their  most  profitable  treatment?  The  ultimate 
purpose  of  the  investigation  may  be  for  a  compari- 
son of  fire  and  wet  methods,  or  for  determining  the 
commercial  value  of  an  ore.  It  is  evident  that  the 
sum  of  the  assay  values  will  not  answer  this  last 
question.  There  are  many  factors  which  enter  into 
the  final  smelting  cost  of  an  ore,  the  chief  one  being 
its  metallurgical  behavior.  In  general  it  is  desired 
to  smelt  an  ore  with  the  least  possible  addition  of 
fluxes,  as  is  compatible  with  the  greatest  net  profit. 
Fluxes  not  only  cost  money  but  also  take  up  valuable 
space  in  the  furnace,  reducing  its  capacity  for  the 
profit-yielding  portion  of  the  charge  and  thus  in- 
creasing smelting  costs.  When  fluxes  contain  metals 
of  value  and  are  subject  to  treatment  charges,  this 
fact  modifies  the  above  statement,  but  then  the  flux 
would  rightly  be  classed  as  an  ore.  On  the  other, 
hand  there  are  disadvantages  resulting  from  insuffi- 
cient fluxing.  The  idea  is  to  use  the  minimum 
amount  of  flux  that  is  allowable  in  view  of  the  in- 
creased cost  of  coke  and  lower  furnace  capacity  with 
highly  silicious  and  limy  slags.  The  increased  loss 
of  metal  from  imperfect  separation  due  to  an  excess 
of  the  basic  oxides  in  slags  of  high  specific  gravity 
must  also  be  reckoned  with.  In  the  case  of  ores 


TYPES  OF  LEAD  SLAGS  165 

containing  undesirable  substances  such  as  A1203, 
MgO,  BaO,  BaS04,  and  ZnO  or  ZnS,  it  must  often  be 
considered  to  what  extent  the  charge  will  have  to 
be  diluted  with  low-grade  or  unprofitable  ores  and 
fluxes  in  order  to  bring  the  mixture  within  smelting 
range.  The  function  of  the  following  tables  and  tab- 
ulated information  is  to  convey  to  the  student  the 
limits  to  which  he  may  go  with  a  slag  when  seek- 
ing to  estimate  the  greatest  probable  net  profit,  and 
also  to  show  the  effect  of  an  increase  or  decrease  of  a 
charge  constituent  upon  a  metallurgical  result. 

LEAD  SLAGS 

In  lead  smelting  the  operator  is  confined  in  the 
beginning  to  the  selection  of  charges  by  the  compara- 
tively rigid  limits  of  the  regular  type  slags,  a  num- 
ber of  which  are  given  in  the  following  tables.  By 
this  statement  is  not  meant  that  the  so-called  'type 
slags'  are  the  only  ones  that  will  work  properly  but 
rather  that  certain  slags  which  have  been  tried 
are  known  to  be  satisfactory.  The  metallurgist  may 
often  have  to  sacrifice  what  he  considers  good  metal- 
lurgical practice  to  obtain  the  greatest  net  profit, 
hence  the  type  slags  are  not  always  followed.  In 
the  following  table  the  analyses  of  type  slags  are 
given  with  the  remarks  on  each.  Under  the  columns 
headed  MnO,  MgO,  BaO,  etc.,  are  given  the  maximum 
percentages  of  these  substances  that  have  been  used 
with  the  corresponding  type.  By  this  it  is  not  in- 
tended to  mean  that  the  entire  list  of  impurities 
would  be  allowable  in  one  slag  to  the  extent  of  the 
maximum  figures  given.  For  the  maximum  percen- 


166 


TESTING  FOR  PROCESSES 


tages  of  one  impurity  in  the  presence  of  others,  the 
reader  is  referred  to  the  explanations  following  the 
table.  A  slag  is  designated  as  whole,  half,  quarter, 
etc.,  according  to  the  ratio  of  the  FeO  +  MnO  to 
the  CaO  +  MgO  +  BaO.  Thus  in  a  half-slag  the  CaO, 
etc.,  is  approximately  one-half  the  percentage  of 
FeO.  When  the  MgO,  etc.,  is  added  to  the  CaO  it  is 
added  according  to  its  replacement  value ;  thus  40.3 
Ib.  of  MgO  is  equivalent  in  fluxing  power  to  56.1 
Ib.  of  CaO.  By  the  word  'charge'  as  in  'per  cent  lead 
on  the  charge, '  is  intended  the  blast-furnace  charge 
including  ore  and  fluxes  but  not  the  coke  or  slag  for 
re-smelting.  In  the  'Table  of  Limits'  is  given  the 
maximum,  minimum  and  average  figures  for  slag 
constituents  found  in  practice,  also  figures  for  other 
data  important  to  silver  lead  smelting.  Unusual  fig- 
ures are  enclosed  in  brackets. 


SLAG  TYPES 


Per  Cent. 


Limits   with   Corresponding 
Type  Slag.     Per  Cent. 


A  ... 

B  ... 

C  ... 

D  .  .  . 

E  ... 

F  . 


rr 

K 

P9 

OS 

0 

O        .  .« 

j 

S?     ' 

O 

5 

28 

28 

6 

31 

23 

4 

34 

24 

35 
31 


30 


32 
30 


1  28 

2  28-30 


38 
38 


40 


47 
48 


50 
54 


17 
21 


20 


11 
12 


I 

~, 

H 
i 

I 

1     ; 

>  • 

5 

2 

-5          1 

0        6 

-10 

5 

2-5          10        6-10 

2-5 


10 


2-5      8-15 
2-5  8 


PERCENTAGE  OF  SLAG  CONSTITUENTS 


167 


REMARKS 

A — Useful  in  case  of  an  excess  of  silicious  ore  and  for  ores 
containing  much  alumina.  It  is  not  adapted  to  ores  high  in 
zinc. 

B — More  fusible  and  faster  running  than  A,  but  still  too 
silicious  for  ores  high  in  zinc. 

C — This  type  may  be  used  successfully  to  work  off  zincy 
ores. 

D — An  excellent  slag  adapted  to  working  irony  ores;  also 
good  for  zincy  slags. 

E — A  good  slag  for  working  off  irony  ores,  where  they  are 
In  excess  or  most  profitable  to  treat. 

F — Not  as  good  as  the  preceding  types  and  of  rather  high 
specific  gravity.  Only  used  in  case  of  an  excess  of  irony 
ores  and  where  silicious  fluxes  are  not  available  at  a  reason- 
able cost. 

SLAG  CONSTITUENT 

Per  cent. 
Maximum.  Minimum.  Average. 


Silica,  SiO2    

Iron  and  manganese,  FeO  +  I\lnO 
Lime  and  magnesia,  CaO  +  MgO 

Manganese,  MnO    

Magnesium    oxide,    MgO 

Barium    oxide,    BaO 

Zinc    oxide,    ZnO 

2Alumina,   A12O3    

Sulphur,  S    

8Lead,  Pb    

"Silver,    Ag    

"Gold,  Au    


36 
54 
28 
43 

5 

8 

15 
10 

1.5 

1.5 

2  oz. 

0.01  oz. 


28 
28 

6 

0 

0 

0 

0 

0 

0.8    ' 

0.5 

0.5  oz. 

Trace 


30-34 

30-45 

12-24 
8-25 
3-  4 
0-  5 
0-  5 
0-  7 
1.0 

0.8-1.0 


GENERAL  DATA 

*Zinc,  volatile    20  10                

Sulphur,  volatile    30  20  2) 

8Fuel   (coke)    17(22)  13  14 

Lead  in   matte    18  7  12-14 

Copper  plus  iron   in  matte 60  55                

Copper   in   matte    0  16  8-12 

Sulphur  in   matte    23  20               

'Matte  fall    20    (4-7)6-8  8-10 

Scrap  iron   on   charge 10  5                

Assay  base  bullion — 

Silver,   Ag    300  oz.          90  oz 

•  Gold,    Au    0.1  oz. 

Lead  on  charge    20  6  10 

Recovery   of   lead 95  90               

Recovery  of  gold    98  .  .                

Recovery  of  silver    98  95               

Successful  runs  have  been  made  with  43%  SiO2. 

28.5%  is  too  high  for  speed  in  smelting  with  a  slag  con- 
taining SiO2  32,  Fe  +  Mn  25%. 

8An  actual  slag  contained:  Pb,  0.8%;  Ag,  1.0  oz.;  Au,  0.004 
oz. 

*Depending  upon  completeness  of  roast. 

"Figured  on  the  basis  of  good  firm  coke  with  10%  ash. 

•Ratio  of  matte  to  ore  plus  flux. 


168  TESTING  FOR  PROCESSES 

EFFECT  OF  SLAG  CONSTITUENTS 

Alumina. — Aluminous  slags  will  generally  be  slow 
running.  They  require  a  higher  heat  to  make  them 
flow  properly  than  their  temperatures  of  formation 
would  indicate.  lies  recommends  that  the  iron  con- 
tent be  kept  well  advanced  and  the  silica  be  between 
29  and  31  per  cent. 

Baryta. — This  enters  both  the  slag  and  matte  and 
increases  the  specific  gravity  of  slag  so  as  to  make 
proper  settlement  of  matte  difficult,  otherwise  a  good 
alkaline  earth  flux.  Baryta  ores  have  been  success- 
fully handled  by  using  a  slag  running  Si02,  35  to 
37%  ;  Fe,  etc.,  20 ;  lime,  27 ;  BaO,  14. 

Lime. — In  general  high  lime  makes  slags  low  in 
lead  except  where  zinc  conflicts  and  requires  the 
iron  to  be  increased. 

Magnesia. — Magnesia  is  especially  troublesome  in 
the  presence  of  zinc  and  alumina,  forming  streaky, 
pasty  slags  high  in  lead.  In  a  slag  containing  8% 
zinc  and  2  to  3%  baryta,  MgO  is  troublesome  and 
5%  will  give  serious  trouble.  Magnesia  alone  and 
under  5  to  8%  will  give  little  trouble  when  intelli- 
gently handled.  It  has  1.4  times  the  replacement 
power  of  lime. 

Potash  and  Soda. — These  elements  in  excess  will 
cause  slags  to  run  high  in  lead  but  good  results  may 
be  obtained  by  using  a  hot  slag  containing  17  to 
22%  of  alkaline  earths. 

Zinc. — Zinc  decreases  the  fluidity  of  the  slag  and, 
while  it  lowers  the  temperature  of  formation,  more 
fuel  is  required  to  make  the  slag  flow  properly  and 


EFFECT  OF  ZINC 


169 


keep  the  lead  content  down.  Zinc  enters  the  matte 
and  decreases  the  specific  gravity  making  it  difficult 
to  settle  properly.  An  excess  of  zinc  is  best  taken 
care  of  by  irony  slags  as  indicated  in  the  table.  It 
is  customary  to  preserve  the  slag  type  by  replacing 
lime  with  one-half  the  zinc  oxide  present,  and  ad- 
justing all  the  constituents  so  as  to  add  up  to  the 
percentage  of  the  original  total.  The  amount  of 
zinc  volatilized  depends  upon  how  well  the  ore  has 
been  roasted  and  the  zinc  changed  into  oxides;  a 
poorly  roasted  ore  will  result  in  a  low  volatilization 
of  zinc.  The  amount  of  zinc  volatilized  and  going 
into  the  matte  will  run  from  10  to  20%.  Following 
is  example  of  slags,  etc.,  resulting  from  smelting  zinc 
retort  residues : 


Base  bullion. 

Ag    87.5    oz. 

Au    ,         .  1.49  oz. 


Slag. 

Per  cent. 

SiO2 

31.1 

FeO 

37  5 

MnO 

1.5 

CaO 

14  1 

ZnO 

10.0 

Pb   .. 

0.77 

Ag,  oz 

1.26 

Matte. 

Ag    19.0    oz. 

Au  0.04  oz. 

Pb    8.7% 

Cu    1.5% 


FURNACE   DATA 

Per  cent.  Saving, 

Lead  on  charge 9.1  percent. 

Coke  on  charge 13.0    Silver    90.8 

Slag  on  charge 32.6    Gold   92.4 

Sulphur  on  charge 3.7   Lead    92.0 

Matte  formed   7.2 

GENERAL  REMARKS 
The  slags  to  be  re-smelted,  consisting  regularly  of 


170  TESTING  FOR  PROCESSES 

the  shells  from  the  slag  dump-pots  and  any  foul  slag 
from  clean-ups  or  unfavorable  runs,  will  average 
about  20  to  30%  of  the  total  'charge,'  and  even  35 
to  40%;  the  term  *  charge'  referring  to  ore  fluxes, 
but  not  the  coke  or  slag  for  re-smelting. 

Iron. — Where  iron  fluxes  are  dear  a  good  slag 
would  be:  Si02,  32%;  Fe,  32;  CaO,  22;  Zn  + 
A1208,  10. 

Base  Bullion. — There  should  be  enough  precious 
metal  on  the  charge  so  that  the  silver  content  of  the 
base  bullion  will  run  between  100  and  300  oz.  per 
ton  and  at  least  0.10  oz.  of  gold  per  ton,  as  the  re- 
finers pay  only  for  0.05  oz.  Au  or  over.  The  richer 
the  bullion  the  greater  the  amount  of  precious  metal 
that  will  be  found  in  the  slag  and  matte.  A  typical 
base  bullion  contained,  Ag  266  oz.,  Au  3.49  oz.,  Pb 
95.0%,  As  0.28,  Cu  0.70. 

Lead. — For  efficient  collecting  of  the  values,  the 
percentage  of  lead  on  the  charge  should  not  be  under 
6  and  better  10.  Over  20%  is  liable  to  force  too  much 
lead  into  the  slag,  though  in  one  smelting  works  as 
high  as  30  to  35%  lead  has  been  placed  on  the  charge 
and  the  resulting  slags  have  contained  less  than  1% 
lead  under  favorable  conditions. 

Fuel. — The  quantity  of  fuel  is  lowered  by  high 
sulphur  and  a  high  percentage  of  lead  on  the  charge, 
also  by  an  easy  running  slag  having  a  low  temper- 
ature of  formation.  The  maximum  fuel  under  these 
conditions  has  been  as  low  as  9  to  10%  of  coke. 

Matte  Fall. — The  quantity  of  matte  formed  will 
depend  upon  the  amount  of  sulphur  available  after 


THE  MATTE  FALL  171 

volatilization  and  slagging.  The  slag  will  contain 
about  1%  sulphur  and  the  volatilization  will  be  from 
20  to  30%  depending  upon  the  condition  of  the  ore ; 
an  unroasted  ore  will  give  a  higher  loss  of  sulphur 
than  a  roasted  product.  The  matte  fall  should  aver- 
age about  10%  in  order  to  keep  down  the  silver-lead 
content  of  the  slag,  and  allow  the  handling  of  more 
impure  and  refractory  ores  than  the  lower  matte  falls 
given  in  the  table.  Low  matte  fall  may  be  used  with 
pure  silicious  ores  in  the  absence  of  BaS04.  It  will 
also  depend  upon  the  cost  of  roasting  and  re-smelt- 
ing the  matte.  Where  manganese  is  substituted  for 
iron  in  the  charge  it  lessens  the  production  of  matte. 
Matte  Content. — The  matte  should  ordinarily  con- 
tain not  less  than  5  and  not  more  than  14%  copper 
to  effectually  cleanse  the  slag  of  silver.  When  in 
excess  of  14  some  of  the  copper  is  reduced  and  goes 
into  the  base  bullion  where  it  must  be  removed  by 
dressing  and  returned  to  the  blast-furnace.  Since 
the  matte  is  roasted  and  returned  to  the  furnace, 
the  copper  accumulates  and  when  in  excess  of  the 
limit  desired,  it  is  removed  and  smelted  separately 
to  form  a  converting  grade  of  copper  matte.  The 
lead,  which  is  mostly  volatilized  in  this  process,  is 
collected  in  the  bag  house.  The  following  charge 
may  be  used  for  the  concentration :  (1)  matte  roasted 
to  4%  sulphur;  (2)  silicious  ores  as  free  from  S,  Au, 
and  Ag  as  possible;  (3)  copper  ores;  (4)  limestone; 
(5)  slag;  (6)  coke,  13  to  14%.  The  slag  should  run 
between  35  and  42%  Si02.  The  matte  formed  should 
contain  about  50%  Cu  and  15  to  20%  Pb.  Scrap 


172  TESTING  FOR  PROCESSES 

iron  may  be  added  to  the  charge  to  the  extent  of  5 
to  10%  to  lower  the  lead  content  of  the  matte,  which 
should  ordinarily  run  between  12  and  14%  but  can 
be  reduced  to  7  or  8%  should  the  conditions  war- 
rant. It  is  not  advisable  to  use  scrap  iron  when 
there  is  much  zincblende  present,  because  metallic 
zinc  is  reduced  and  oxidizes  to  ZnO  in  the  upper 
part  of  the  furnace,  forming  accretions  and  pre- 
venting free  combustion  of  the  coke  from  the  coating 
of  oxide  upon  it.  A  lack  of  fuel  to  sufficiently  super- 
heat the  slag  and  give  proper  reducing  conditions  is 
quite  apt  to  produce  a  matte  high  in  lead.  In  gen- 
eral a  slag  of  25  to  28%  FeO  is  best  suited  for  pro- 
ducing a  matte  low  in  lead.  The  lead  also  increases 
with  the  copper  in  the  matte,  namely,  a  40%  copper 
content  may  call  for  20%  lead.  A  typical  matte  anal- 
ysis is:  S,  20.29%;  Fe,  35.0;  BaO,  0.73;  Zn,  6.42; 
Pb,  10.96 ;  Cu,  14.8 ;  Ag,  77  oz.  per  ton ;  Au,  0.18  oz. 
per  ton;  specific  gravity,  4.64. 

Speiss. — When  sufficient  arsenic  is  present  to  form 
a  speiss,  2.3  times  the  weight  of  arsenic  equals  the 
weight  of  iron  needed  to  form  it.  This  must  be  sub- 
tracted from  the  total  iron  in  the  charge  as  it  is  not 
available  for  slag  or  matte.  The  per  cent  of  speiss 
formed  may  be  approximated  by  running  a  regular 
lead  assay  on  the  ore  and  noting  the  weight  of  the 
speiss  button  found  with  the  lead.  Ores  forming 
much  speiss  should  be  run  with  a  silicious  slag  in 
order  that  the  temperature  of  the  furnace  may  be 
high  enough  to  keep  the  speiss  liquid. 

Losses. — In  best  practice  90  to  95%  recovery  of 


LEAD  ROASTING  173 

the  lead  and  98%  of  the  gold  and  silver  can  be  ob- 
tained. 

Flue  Dust. — The  amount  of  flue  dust  will  run  from 
0.3  to  3.5%  (and  more)  of  the  total  charge  accord- 
ing to  the  amount  of  fine  material  and  the  strength 
of  the  blast.  An  average  figure  would  be  2x/2  P®r 
cent. 

Lead  Roasting. — The  question  when  to  roast  will 
be  decided  by  the  matte  fall  desired.  Sulphur  is 
commonly  debited  at  25c.  per  unit,  which  represents 
the  approximate  cost  of  handling  and  reworking  the 
resulting  matte.  Then,  since  for  each  unit  of  sul- 
phur driven  off  in  a  roast,  25c.  is  saved,  the  advis- 
ability of  roasting  may  be  estimated  by  taking  into 
account  the  cost  of  roasting  and  the  resulting  losses 
which  are  from  2  to  1%  of  the  lead  and  1  to  5%  sil- 
ver, but  can  be  kept  within  3  to  4%  lead  and  1  to  2% 
silver  with  a  trace  of  gold.  The  change  in  weight  of 
a  roast  approximates  the  replacement  of  sulphur 
by  oxygen  and  also  in  the  case  of  a  pot  roast  the 
loss  of  C02  of  the  limestone  when  used. 

COPPER  SLAGS 

The  copper  slags  are  usually  classed  according  to 
the  ratio  of  the  oxygen  on  the  base  side  to  the  oxy- 
gen on  the  acid  side.  Iron  and  lime  oxides  being 
the  chief  bases  and  silica  the  principal  acid. 

(1)  Subsilicate    3ROSiO2 

Ratio  3  to  2. 

( 2 )  Singulosilicate    2ROSiO2 

Ratio  2  to  2. 


174  TESTING  FOR  PROCESSES 

(3)  Sesquisilicate    4RO3SiO2 

Ratio  2  to  3. 

(4)  Bisilicate ROSiO2 

Ratio  1  to  2. 

(5)  Trisilicate   2RO3SiO2 

Ratio  1  to  3. 

In  these  equations  EO  represents  the  oxide  of  any 
base  or  combination  of  bases. 

The  silicate  degree  of  a  complex  slag  is  found  by 
taking  the  percentage  of  oxygen  in  the  Si02  or  acid 
side  and  that  on  the  base  side  by  the  use  of  the  fol- 
lowing table  and  determining  the  ratio. 

Per  cent 
oxygen. 

Silica,   SiO2    53.35 

Iron  oxide,  PeO 22.2 

Manganese  oxide,  MnO   22.6 

Lime,  CaO   28.6 

Magnesia,  MgO  40.0 

Baryta,   BaO    10.5 

Alumina,  A12O3  47.0 

For  example,  take  slag  of  the  following  analysis : 

Per  cent  O     Per  cent  O 
Per  cent,    on  base  side,  on  acid  side. 

Si03    40.0  ...  21.3 

PeO    30.0  6.7 

MnO    5.0  1.1 

CaO    20.0  5.7 

MgO   t . . . .         5.0  2.0 

15.5  =  x       21.3  =  y 
x:y  as  15.5:21.3  =  1:1.37;  approximately  3:4  or  3RO,2SiO2. 

The  subsilicate  would  only  be  used  in  case  of  a 
great  excess  of  basic  ores,  usually  irony,  with  sili- 


SELECTION  OP  A  SLAG  175 

cious  fluxes  not  available  at  a  reasonable  cost.  This 
type  requires  a  rather  high  temperature  to  run  prop- 
erly ;  they  are  very  corrosive  to  crucible  linings,  and 
are  of  too  high  specific  gravity  to  make  a  clean  slag, 
and  hence  are  generally  unsuited  to  economical 
smelting  except  in  extreme  cases.  The  singulo  sili- 
cate enters  the  range  of  ordinary  commercial  slags 
and  is  used  where  iron  or  other  bases  are  in  excess. 
It  is  still  of  too  high  specific  gravity  to  give  a  per- 
fect settlement  of  the  matte  and  is  seldom  made 
where  silicious  ores  are  obtainable  at  a  fair  price. 
The  sesqui  silicate  is  a  mixture  of  the  singulo  silicate 
with  a  bisilicate,  and  in  the  neighborhood  of  this 
silicate  degree  the  majority  of  the  copper  slags  may 
be  placed.  It  has  a  low  melting  temperature  and 
runs  smoothly  without  forming  a  thick  crust  over  its 
surface.  It  is  low  enough  in  specific  gravity  to  per- 
mit a  complete  and  rapid  settling  of  the  matte.  It  is 
perhaps  the  most  desirable  slag  when  conditions  per- 
mit its  use.  The  bisilicates  are  used  in  case  of  an 
excess  of  silicious  ores.  While  they  require  a  higher 
percentage  of  coke  on  the  charge  than  the  preceding 
slags,  still  when  they  are  properly  melted,  give  a 
slag  of  low  specific  gravity  and  a  very  clean  separa- 
tion of  the  matte.  A  trisilicate  comes  more  within 
the  range  of  the  iron  blast-furnace.  The  following 
table  gives  the  limits  of  the  different  slag  constit- 
uents as  found  in  modern  practice. 


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TYPES  OF  SMELTING  177 

By  'ordinary  matte  smelting'  is  meant  that  prac- 
tice which  uses  raw  or  roasted  ore  with  low  concen- 
tration, of  say  4  or  5  into  1,  and  10  to  15%  of  fuel, 
and  primarily  for  the  reduction  of  copper  ores  with 
gold  and  silver  as  by-products.  By  '  semi-pyrite ' 
smelting  is  meant  that  type  which  uses  a  high  concen- 
tration ratio,  silicious  slags,  as  small  amount  of  cop- 
per on  the  charge  as  possible,  pyrite  being  used  as  a 
flux  and  fuel  if  available,  but  primarily  smelting  is 
for  the  collection  of  gold  and  silver  from  unroasted 
ores  that  are  not  suitable  for  amalgamation  or  leach- 
ing methods.  By  'pyrite  smelting'  is  meant  a  smelt- 
ing of  raw  heavy  sulphide  ores  for  the  collection  of 
their  copper  content,  also  gold  and  silver,  with  the 
least  possible  addition  of  lime  and  silicious  flux.  It 
is  not  possible  to  sharply  define  all  types  of  smelting 
by  these  three  divisions  as  there  will  be  gradations 
of  one  into  the  other  as  conditions  require.  Ores  for 
regular  pyrite  smelting  should  contain  60%  of  py- 
rite or  its  equivalent. 

Example. — The  use  of  the  table  may  be  illustrated 
as  follows.  In  deciding  on  a  slag  for  smelting  a  cop- 
per ore  by  ordinary  matte  smelting  it  is  desirable  to 
see  how  high  an  alumina  content  may  be  used.  Op- 
posite A1203  in  the  column  headed  'Constituents  of 
Slags'  and  in  the  column  headed  'Ordinary  Matte 
Smelting,'  'Average  %  Max.'  (maximum)  is  found 
the  figure  10  which  indicates  that  under  average  con- 
ditions 10%  A12O3  may  be  used  in  a  slag.  In  a  like 
manner  under  the  column  headed  '  Special  %  Max. ', 
is  the  figure  15  which  indicates  that  in  special  cases 


178  TESTING  FOR  PROCESSES 

15%  A12O3  may  be  used  in  a  slag.  These  figures  are 
supplemented  by  the  remarks  on  alumina  which  are 
given  following  the  table.  The  table  may  be  used 
in  a  like  manner  to  find  the  average  figures  for  other 
slag  constituents  for  all  common  types  of  smelting, 
also  other  data  as  matte  fall,  fuel  on  charge,  etc. 
The  column  heading  'Min.',  is  intended  as  an  abbre- 
viation for  minimum.  These  columns  contain  the 
minimum  or  lowest  figures  found  in  average  as  well 
as  special  cases. 

GENERAL  REMARKS 

Arsenic  and  Antimony. — These  constituents  are 
volatilized  to  a  large  extent  in  pyrite  or  semi-py- 
rite  smelting.  They  may  be  almost  entirely  elimin- 
ated in  reconcentrating  the  matte. 

Alumina. — The  following  deductions  may  be  made 
from  Hoffman's  investigations  of  alumina  replace- 
ment : 

Alumina  cannot  successfully  replace  Si02,  hence 
with  much  A1203  present  the  SiO2  must  be  kept  high. 
Alumina  cannot  replace  FeO  to  any  extent  with  an 
undue  rise  in  the  temperature  of  formation  of  the 
slag.  It  is  within  reason  to  figure  A1203  as  replac- 
ing CaO.  Taking  all  of  the  above  into  consideration 
shows  that  the  formation  temperature  of  the  slag 
may  be  kept  unchanged  if  alumina  were  figured  as 
replacing  part  of  the  CaO  and  Si02.  This  method 
found  practical  application  in  smelting  the  alum- 
inous Cripple  Creek  ores  in  silver-lead  blast-fur- 
naces. Although  the  temperature  of  formation  of 
aluminous  slags  is  low,  yet  they  require  much  super- 


EFFECT  OF  BARYTA  179 

heat  to  make  them  flow  properly.  In  one  case  with 
slags  between  a  sesqui  silicate  and  a  bisilicate,  alu- 
mina was  satisfactorily  figured  as  a  base.  In  another 
case  with  a  higher  percentage  of  alumina  it  was 
more  satisfactorily  figured  as  having  two-thirds  the 
replacement  power  of  silica. 

Baryta. — The  same  remarks  apply  as  given  under 
lead  slags.  The  Tyee,  British  Columbia,  plant  ran 
slags  which  contained:  Si02,  34.7% ;  FeO,  17.7;  BaO, 
28.5.  A  still  higher  percentage  might  be  used  in  ex- 
ceptional cases. 

Heavy  Spar  (BaS04). — This  constituent  is  partly 
reduced  to  BaS  and  enters  the  slag  and  matte,  de- 
creasing their  difference  in  specific  gravity,  besides 
making  thick  infusible  slags  and  tending  to  form 
wall  accretions. 

Galena. — The  lead  tends  to  volatilize  and  carry  sil- 
ver with  it,  and  is  an  unwelcome  addition  to  the 
charge  on  this  account. 

Iron  and  Manganese. — Their  behavior  is  about  the 
same,  and  MnO  may  be  safely  figured  as  FeO  up  to 
28  to  30%  MnO. 

The  following  is  an  example  of  a  slag  having  high 
SiO2  and  CaO  and  low  FeO ;  reported  as  fluid,  flow- 
ing readily  but  chilling  quickly.  It  was  used  at  the 
Rapid  City,  South  Dakota,  smelter. 

Per  cent. 

SiO2    50.20 

FeO    16.35 

CaO 28.30 

A12O3    4.20 

Total    .  99.05 


180  TESTING  FOR  PROCESSES 

When  FeO  alone  is  available  for  a  base  the  fol- 
lowing slag  might  be  used : 

Blast-furnace.  Reverberatory. 
Per  cent.          Per  cent. 

Si02    38  38-42 

FeO    62  62-58 

Lime  and  Magnesia. — When  using  the  maximum 
amounts  of  MgO,  the  slag  is  much  less  fluid  than  the 
temperature  of  formation  would  indicate  and  re- 
quires careful  handling.  It  is  essential  while  using 
high  magnesia  slags  to  keep  up  the  percentage  of 
iron  to  at  least  20.  With  15  to  IS%  FeO  the  issue 
will  be  doubtful.  ZnO  and  MgO  form  a  troublesome 
combination.  The  effect  of  replacing  CaO  with  ZnO 
is  to  lower  the  temperature  of  formation,  but  the  re- 
sultant slags  show  so  little  fluidity  that  much  super- 
heat is  necessary  for  them  to  flow  properly.  When 
within  reasonable  limits,  as  given,  MgO  is  a  good 
flux  as  it  has  1.4  times  the  fluxing  power  of  CaO.  The 
effect  of  high  lime  slags  on  formation  temperatures 
is  clearly  shown  in  the  table  of  slags  following.  The 
effect  of  lime  is  to  lessen  the  specific  gravity  of  the 
slags,  which  is  productive  of  more  perfect  matte  set- 
tlement, providing  the  slag  has  a  sufficiently  low  tem- 
perature formation  so  that  it  may  receive  enough 
superheat  to  flow  properly. 

Sulphur. — The  volatilization  depends  upon  the  na- 
ture of  the  ore.  Under  similar  conditions  a  roasted 
ore  will  give  a  lower  sulphur  loss  than  an  unroasted 
one.  The  conditions  under  which  the  furnace  is  run 
also  governs  the  loss  of  sulphur ;  with  high  fuel,  giv- 
ing a  strong  reducing  effect,  less  volatilization  is  to 


MATTE  AND  CONCENTRATION  RATIO  181 

be  expected  than  with  low  fuel  and  a  large  volume 
of  blast. 

Zinc. — The  zinc  entering  the  slag  will  depend  upon 
the  amount  volatilized  and  going  into  the  matte.  The 
same  remarks  apply  as  given  under  zinc  in  lead  slags. 
Under  the  roasting  conditions  for  copper  ores  much 
of  the  ZnS  is  left  undecomposed  and  enters  the  slag. 
Since  zinc  is  volatile  it  causes  a  loss  of  silver  by  vola- 
tilization, forms  wall  accretions,  condenses  with 
the  flue  dust,  and  is  thus  returned  to  the  furnace. 
When  it  is  present  in  the  charge  to  the  extent  of  1.5 
to  2%  the  bad  effects  are  noticeable,  and  10%  is 
prohibitive. 

Mattes  and  Concentration  Ratios. — It  is  not  often 
desirable  to  use  a  matte  of  less  than  5%  copper,  for 
efficienct  collection  of  values.  In  cases  where  copper 
is  lacking  to  bring  the  matte  up  to  this  standard, 
it  would  be  advisable  to  re-charge  enough  roasted 
matte  to  reduce  this,  providing  the  additional  sav- 
ing paid  for  the  cost  of  the  operations.  The  cop- 
per in  the  matte  should  be  proportional  to  the  matte 
fall.  With  high  concentration,  10%  copper,  with 
low  concentration  2  to  3%  copper  is  sufficient.  The 
grade  of  the  matte  should  also  vary  directly  with 
the  amount  of  gold  and  silver  and  impurities  pres- 
ent. In  general,  it  is  not  good  practice  to  use  high 
concentration  ratios  such  as  18  to  22  tons  of  charge 
into  1  ton  of  matte,  except  where  conditions  de- 
mand this  or  with  exceptionally  pure  ores  in  fur- 
naces of  200  or  more  tons  capacity.  A  10  into  1  or 
12  into  1  concentration  is  about  the  most  satisfac- 


182  TESTING  FOR  PROCESSES 

tory  limit.  For  example  in  a  furnace  36  by  84  inches, 
making  a  slag  assaying  Si02  34%,  FeO  33,  CaO  20, 
concentration  of  10  into  1  caused  the  furnace  to 
crust  badly  and  often  necessitated  refeeding  of 
matte.  A  15%  matte  fall  proved  more  satisfactory. 

When  the  matte  from  the  first  smelting  does  not 
come  up  to  converting  or  shipping  grade,  say  35% 
copper,  it  may  be  re-concentrated  in  the  same  fur- 
nace, or  if  the  daily  output  is  sufficient,  in  a  separate 
furnace.  The  charge  may  consist  of  large  lumps  of 
raw  matte,  silicious  ores,  limestone,  and  rich  slag. 
The  process  is  inexpensive  as  computed  per  ton  of 
original  ore,  and  serves  to  remove  As,  Sb,  Pb,  and 
Zn  which  are  detrimental  to  refining.  The  grade  of 
matte  produced  for  converting  ranges  ordinarily  be- 
tween 35  and  50%  but  more  often  between  45  and 
50%  copper. 

Peters  says,  "I  would  advise  no  metallurgist  to 
undertake  a  new  enterprise  (semi  pyrite  smelting) 
unless  he  sees  his  way  clear  to  keeping  the  Si02  down 
toward  45%  and  his  FeO  up  to  at  least  20%.  If 
much  A12O3  is  present,  I  would  advise  that  in  his 
preliminary  calculations,  he  lower  the  Si02  in  his 
slag  0.66%  for  each  unit  of  A1203  in  the  mixture 
above  5%.  If  ZnS  above  5%  be  present,  I  would 
allow  an  increase  of  1%  FeO  in  my  slag  for  each  unit 
excess  ZnS  in  my  mixture."  He  further  adds  that 
these  figures  are  purely  empirical,  and  gives  other 
cautions. 

The  size  of  the  blast-furnace  has  a  decided  effect 
upon  the  permissible  limits  of  the  slag  constituents. 


LOSS  OP  WEIGHT  IN  ROASTING  183 

On  account  of  the  small  amount  of  the  slag  and 
matte  flowing  away  from  the  furnaces  of  low  capac- 
ity and  hearth  area,  a  much  more  fusible  and  easy 
running  slag  must  be  used  than  would  be  neces- 
sary in  furnaces  of  large  capacity  used  in  modern 
practice.  Intermittent  tapping  will  allow  the  run- 
ning of  slags  that  could  not  be  handled  with  a  con- 
tinuous flow  into  a  settler.  For  furnaces  of  less  than 
150  tons  daily  capacity,  the  following  limits  should 
be  observed  when  possible :  Si02  30  to  38%  ;  FeO  30 
to  40 ;  CaO  10  to  25.  One  of  the  best  all  round  slags 
has  the  following  composition:  Si02  36%;  FeO  33; 
CaO  21. 

Loss  in  Weight  Due  to  Roasting. — When  an  ore  is 
assumed  to  be  roasted,  as  must  often  be  done  in  pre- 
liminary calculations,  it  is  necessary  to  approximate 
its  analysis  without  resort  to  actual  trial.  Since  oxy- 
gen (16)  is  one-half  the  atomic  weight  of  sulphur 
(32),  in  roasting  from  sulphides  to  oxides  the  loss 
in  weight  can  be  approximated  as  one-half  the  loss 
of  sulphur.  Example : 

Analysis  original  ore.  Per  cent  in  100  Ib.  ore. 

SiO2 60     or   60  Ib. 

Pe    20     "      20  " 

S    20     "      20  " 

Total    100    in    100  Ib. 

Roasted  to  a  product  containing  5%  sulphur:  loss 
in  weight  =  100  -  ~  =  92%  Ib. ;  therefore  in  92% 
Ib.  of  roasted  ore  there  is  Si02  60  Ib.,  Fe  20  Ib. 
The  new  approximated  analysis  would  be : 


184 


TESTING  FOR  PROCESSES 


Per  cent. 


(50 


Si02    92^       X  100  =  65.0 

Fe      QJfc        X  100  =  21.6 
S  assumed  =   5.0 

The  following  slag  tables  (by  Hoffman)  show  the 
variation  of  constituents  possible  in  the  different  sili- 
cate degrees  with  the  attendant  effect  on  the  tem- 
perature formation.  The  metallurgist  must  not  be 
entirely  guided  by  the  low  formation  points  of  the 
highly  silicious  bisilicates  as  they  need  a  tempera- 
ture considerably  higher  to  flow  properly.  They 
may  be  used  successfully  in  semi  pyrite  smelting  on 
account  of  the  high  temperature  attained. 
HOFFMAN'S  SLAG  TABLES 


Subsilicate 

3RO 

,Si02. 

3  to  4 

Silicate 

3RO,2Si02. 

Chemical   composi- 

Melting 

Chemical  composi- 

Melting 

tion 

of  slag. 

point. 

tion 

of  slag. 

point. 

Si02, 

FeO,     CaO, 

Deg. 

SiO2, 

FeO,    CaO, 

Deg. 

% 

% 

% 

C. 

% 

% 

% 

C. 

21.70 

78.30 

0 

1220 

35.70 

64.30 

0 

1140 

21.95 

74.05 

4 

1230 

36.05 

60.00 

4 

1110 

22.20 

69.08 

8 

1220 

36.40 

55.60 

8 

1090 

22.49 

65.51 

12 

1200 

36.80 

51.20 

12 

1070 

22.70 

61.30 

16 

1240 

37.30 

46.70 

16 

1090 

22.95 

57.05 

20 

1250 

37.75 

42.25 

20 

1110 

23.20 

52.80 

24 

1210 

38.16 

37.84 

24 

1130 

23.45 

48.55 

28 

1190 

38.56 

33.44 

28 

1150 

23.70 

44.30 

32 

1170 

38.95 

29.04 

32 

1160 

23.94 

40.06 

36 

1170 

39.37 

24.63 

36 

1170 

24.20 

35.80 

40 

1230 

39.78 

20.22 

40 

1190 

24.45 

31.55 

44 

1310 

40.20 

15.80 

44 

1290 

24.48 

27.52 

48 

1430+ 

40.60 

11.40 

48 

1430+ 

24.95 

23.05 

52 

1430+ 

41.02 

6.98 

52 



HOFFMAN'S  SLAG  TABLE 


185 


Singulosilicate  2RO,SiO2. 

Sesquisilicate  4RO,3SiO2. 

Chemical  composi-     Melting 

Chemical  composi-     Melting 

tion  of  slag.             point. 

tion  of  slag.             point. 

SiO,,,         FeO,     CaO,     Deg. 

SiO2,         FeO,     CaO,     Deg. 

%            %         %        C. 

%            %         %        C. 

29.20         70.80         0       1270 

38.46         61.54         0       1120 

29.75         66.25         4       1250 

38.90         57.10         4       1090 

30.09         61.91         8       1240 

39.34         52.66         8       1060 

30.42         57.58       12       1220 

39.78         48.22       12       1060 

30.76         53.24       16       1170 

40.22         43.78       16       1090 

31.07         48.90       20       1205 

40.66         39.34       20       1130- 

31.40         44.60       24       1190 

41.11         34.89       24       1150 

31.70         40.30       28       1170 

41.54         30.86       28       1160 

32.10         35.90       32       1150 

41.99         26.01       32       1165 

32.30         31.70       36       1130 

42.42         21.58       36       1190 

32.70         27.30       40       1150 

42.87         17.13       40       1250 

33.10         22.90       44       1190 

43.31         12.69       44       1330+ 

33.44         18.56       48       1270 

43.75          8.26       48       .... 

33.79         14.21       52       1430+ 

44.19           3.81       52       

Bisilicate  RO,SiO2. 

Cross  series,  FeO  :  CaO  =  2:1. 

45.45         54.55         0       1110 

18.67        54.23     27.10     1190 

46.00         50.00         4       1170 

25.61        49.60     24.79     1180 

46.53         45.47         8       1030 

31.47        45.68     22.85     1190 

47.04         40.96       12       1050 

36.47        42.36     21.17     1180 

47.56         36.44       16       1090 

40.80        39.46     19.74     1160 

48.02         31.98       20       1130 

44.55        36.97     18.48     1140 

48.57         27.43       24       1170 

47.86        34.77     17.37     1120 

49.19         22.81       28       1200 

50.82        32.78     16.40     1115 

49.60         18.40       32       1250 

53.44        31.04     15.52     1110 

50.11         13.89       36       1330 

55.81        29.46     14.73     1110 

50.63           9.37       40       1430 

57.95        28.04     14.01     1130 

51.14          4.86       44       .... 

59.87        26.75     13.38     1310+ 

51.65          0.35       48       

51.73          0.00       52       

—  — 

186  TESTING  FOR  PROCESSES 

CALCULATION  OF  A  SILVER-LEAD  BLAST- 
FURNACE  CHARGE 

Conditions. — Assume  that  it  is  desired  to  smelt  as 
much  dry  silicious  ore  as  is  possible  with  only  enough 
lead  for  a  collector  and  that  zinc  sulphide  is  pres- 
ent in  the  lead  ore  to  an  extent  which  would  lead 
to  choice  of  a  slag  of  the  C  type,  which  is  especially 
adapted  to  working  zincy  ores  and  at  the  same  time 
is  as  silicious  as  possible.  It  will  be  necessary  to  put 
enough  roasted  matte  in  the  charge  so  that  this  prod- 
uct will  not  accumulate.  Clippings,  foul  slag,  etc., 
will  have  to  be  introduced  after  the  plant  is  run- 
ning regularly.  The  charge  may  be  started  with  50 
Ib.  roasted  matte,  figuring  the  whole  charge  on  the 
basis  of  1000-lb.  units  for  convenience.  For  the 
first,  or  starting  charge  the  most  favorable  slag  pos- 
sible would  be  chosen  so  that  the  irregularities  of 
starting  may  be  more  easily  overcome.  The  anal- 
yses of  ore,  etc.,  are  given  on  the  accompanying 
charge  sheet. 

The  next  consideration  is  to  get  a  safe  amount  of 
lead  on  the  charge,  assumed  at  10%.  This  may  be 
done  by  adding  enough  galena  to  furnish  100  Ib.  of 
lead  or  167  Ib.  galena.  Next,  purely  by  guess  and 
later  by  experience,  275  Ib.  silicious  ore  is  put  in 
as  the  probable  amount  required  to  make  a  charge 
of  1000  Ib.  The  next  step  is  to  figure  out  the  weights 
of  the  various  constituents  Si02,  CaO,  etc.,  The  to- 
tal weight  of  the  Si02  from  the  galena,  matte,  and 
silicious  ore,  amounts  to  about  220  Ib.  This  gives 


RATIO  OP  IRON  AND  LIMESTONE  TO  SILICA    187 

data  enough  to  make  the  first  estimate  of  the 
amounts  of  limestone  and  iron  ore  required. 

Factors. — The  Fe  in  the  slag  is*^  or  0.78  times 
the  silica.  Knowing  the  pounds  of  silica  going  into 
the  slag,  the  amount  of  iron  required  to  go  with  it  to 
make  a  slag  of  the  given  analysis,  can  be  found  by 
multiplying  the  weight  of  the  silica  in  the  charge  by 
0.78.  220  X  0.78  =  172  Ib.  Fe  required  or  172  -5-  0.60 
=  287  Ib.  of  iron  ore  since  the  latter  only  contains 
60%  iron.  Since  the  matte  will  require  iron  and  both 
the  iron  ore  and  limestone  contain  silica  which  is 
not  included  in  the  above  amount,  of  220  Ib.,  a 
slightly  larger  quantity  of  iron  ore,  325  Ib.  must  be 
taken.*  220x0.5  —  110  Ib.  CaO  or  120-^0.54  = 
200  Ib.  of  limestone  or,  allowing  for  the  extra  silica, 
about  225  Ib.  would  be  required.  The  analysis  of  the 
coke  ash  must  be  converted  into  terms  of  coke.  The 
coke  has  10%  ash.  The  ash  contains  75%  silica,  then 
the  coke  will  contain  10%  of  the  75  or  7.5  silica. 
Complete  finding  the  weights  of  the  constituents  and 
add  up  the  total  amounts. 

All  the  iron  is  not  available  for  the  slags  as  some 
of  it  goes  into  the  matte.  The  weight  of  the  matte 
must  first  be  computed  from  the  amount  of  sulphur 
available.  The  slag  contains  0.8%  sulphur.  Since 
the  silica  is  34%  of  the  slag  the  weight  will  be  267 
-i-  0.34  =  785  Ib.  and  will  contain  785  times  0.008  = 
6  Ib.  of  sulphur.  Then  allowing  for  20%  volatiliza- 
tion 20%  of  23  =  5  Ib.  of  sulphur  may  be  calculated, 
making  a  total  loss  of  11  Ib.  and  leaving,  23  — 11  = 


*The  amount  of  iron  in  the  incoming  matte  is  25  pounds. 


188  TESTING  FOR  PROCESSES 

12  lb.  of  sulphur  available  for  the  matte.  This 
amount  will  make  12  -f-  0.20  =  60  lb.  matte  con- 
taining 20%  sulphur.  To  find  the  iron  in  the  matte 
the  copper  must  first  be  determined.  Practically  all 
the  copper  goes  into  the  matte  if  sufficient  sulphur 
has  been  provided.  There  is  then  8  lb.  of  copper  in 
59.5  lb.  of  matte  or  (8 -f- 60)  X  100  =  13%  copper. 
Then  the  per  cent  of  iron  is  60  —  12.9  =  47.1%  Fe. 
The  weight  of  the  iron  going  into  the  matte  will  be 
47%  of  60  lb.  or  28  lb.  which  must  be  subtracted 
from  the  total  iron  in  the  charge,  leaving  the  iron 
available  for  the  slag.  If  there  were  any  arsenic 
present  to  form  a  speiss,  2.3  its  weight  must  be  sub- 
tracted from  the  total  iron.  The  next  step  is  to  see 
if  the  iron  and  calcium  oxides  are  present  in  the  right 
amounts  to  form  the  slag  given.  (Computations  C.) 
According  to  the  factors  the  Fe  is  0.78  times  Si02. 
There  is  267  lb.  of  Si02  present  hence  there  should 
be  0.78  times  267  =  209  lb.  of  iron  present,  but  only 
206  lb.  is  available  hence  there  is  a  deficiency  of  3 
lb.  or  in  terms  of  iron  ore,  3-^0.60  =  5  lb.  The 
charge  will  have  to  be  corrected  by  adding  this 
amount  and  if  the  error  is  very  large  all  the  compu- 
tations will  have  to  be  gone  over.  According  to  the 
factor  the  CaO  is  0.5  times  the  Si02  hence  267  times 
0.5  =  134  lb.  CaO  is  needed.  There  is  139  lb.  CaO 
on  the  charge  which  is  an  excess  of  5  lb.  CaO  or  5  -r- 
0.50  =  10  lb.  of  limestone  which  is  added  to  the 
charge. 

If  desirable  the  approximate  amount  of  base  bul- 
lion may  next  be  calculated,  by  subtracting  the 
weight  of  the  lead  in  the  matte  and  slag  from  the 


CALCULATION  OF  BULLION 


189 


total  weight  in  the  charge.  This,  of  course,  does 
not  account  for  other  metallurgical  losses  which  are 
offset  in  a  measure  by  the  impurities  entering  the 
bullion.  Lead  in  the  matte  may  be  determined  as 
follows:  60  Ib.  of  matte  assaying  15%  lead  would 
contain  60  times  0.15  =  9  Ib.  of  lead.  That  in  the 
slag  may  be  found  as  follows :  785  Ib.  of  slag  assay- 
ing 0.8%  would  contain  785  times  0.008  =  6  Ib.  lead. 
Total  9  +  6=  15  Ib.  of  lead  or  107  -  15  ==  92  Ib.  base 
bullion. 

Knowing  the  weight  of  the  base  bullion  its  prob- 
able assay  may  next  be  computed.  First  compute 
the  total  weight  of  the  gold  and  silver  in  the  ores 
(not  including  the  matte)  from  the  assay  values 
(Computation  E),  subtract  from  these  amounts  the 
weights  going  into  the  slag  and  the  result  will  be 
the  amounts  going  into  the  base  bullion,  not  taking 
into  account  the  other  metallurgical  losses  such  as 
volatilization  and  flue  dust.  The  reason  the  matte  is 
not  taken  into  account  is  that,  while  silver  and  gold 

CHARGE  SHEET 


Af*me  of  Or-e. 

«»o 

We 

IVlif 

9M 

Cry 

P 
% 

fa 

sn 

•/„ 

Ji 
wt 

Hi. 
% 

•></  Mn 
Wt. 

c«o<* 

% 

'w/° 

i 
% 

vvt 

c 
•/, 

we 

Rot-sttd  tna.tte.  --- 

SO 

IS 

8 

a 

1 

5o 

iS 

/ 

H.5 

to 

<>1 

/6 

Ib 

10 

/'6 

10 

11 

t 

a. 

Silicious  Ofe  

i7f 

?" 

/?.?, 

V 

/f 

Cor*»e.ti 

Ifon  ofe  4 

0/J-- 

330 
3Z5 

10 

.« 

60 

•t<)5 

C«ft 
Lim-estone.  

ect 

iar>- 

lib 
*15 

3 

7 

5<C 

/a/ 

(/'. 

o) 

5 

8 

in 

1.8 

J 

o.y 

/ 

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3. 

7~otA/s  -  -  --- 

IO& 

% 

*u 

131 

ZZ 

•ft 

a 

190 


TESTING  FOR  PROCESSES 


are  introduced  into  the  charge  by  the  roasted  matte, 
practically  the  same  amounts  will  be  taken  out  by 
the  matte  that  is  formed.  Where  matte  is  fed  ir- 
regularly this  short  cut  cannot  be  taken. 

SLAG  Per  cent. 

SiO2    34.0 

FeO    34.0  (Fe,  26.4%) 

CaO    17.0 

Pb    0.8 

S    0.8 

Ag    0.5  oz. 

COKE  (10%  ash,  1.5%  S) 
Coke  ash.  Per  cent. 

Si02    75 

Fe    18 

CaO    9 

MATTE  Per  cent. 

S    20 

Pb    15 

Cu  -f  Fe  60 

COMPUTATION    (A)    FACTORS 

?^  =0.78  iron  factor. 

o4«0 

1™  =  0.5  lime  factor. 

COMPUTATION    (B> 

^g  X  0.8   =   6  Ib.  S  in  slag. 
22.7  X  0.20  =   5    "     "  volatilized. 

Total     11   "     "  lost. 
.-.23  —  11  =  12    "     "  to  matte. 

A   x  100  =  13%  Cu  in  matte. 

60 

.  • .  60  —  13  =  47%  Fe  in  matte. 
120   =  60  Ib.  of  matte. 

60  X  (0.60  — 0.13)  =28  Ib.  Fe  in  matte. 


ESTIMATION  OF  CHARGES  191 


COMPUTATION    (c) 

267  X  0.78  =  209  Ib.  Fe  needed. 

206  "  "  in  charge. 

3  "  "  lacking. 

3  -T-  0.60  =  5  Ib.  iron  ore  to  be  added. 
267  X  0.5  =  134  Ib.  CaO  needed. 

139  "   "  in  charge. 

5    "      "     excess. 
5  -j-  0.54  =  9  Ib.  limestone  excess. 

COMPUTATION    (D) 

267  x  0.008  =  6  Ib.  Pb  in  slag. 
107  —  (6  +  9)  =92  Ib.  wt.  base  bullion. 
_§?_  x  100  =  5.8%  matte  fall. 

1036 

60  X  0.15  =  9  Ib.  Pb  in  matte. 

COMPUTATION    (E) 

Assays  galena  =  Ag,  50.0  oz. 

16&20oo°'0  =  4'125  °Z'  Ag  in  galena" 
267  x  Q.5    =  0.197    "    "     "    slag. 

0.34  X  2000 

3.928    "     "      "    base  bullion. 
Silicious  ore=:Au,  1.00  oz. 
276  x  IM  =  0.137  oz.  Au  in  silicious  ore. 

2000 

trace         "      "   slag. 


0.137    '  '    base  bullion. 

ASSAY    OF    BASE    BULLION 

2000  X  3-928  =  81.0  oz.  Ag  per  ton. 

92 

2000  X  0.137  _  2  9g   QZ  ton  Au 

92 


192  TESTING  FOR  PROCESSES 

CALCULATIONS  OF  A  COPPER  BLAST- 
FURNACE CHARGE 

The  problem  is  to  select  the  best  charge  to  smelt 
an  ore  containing  Si02,  50% ;  Fe,  15;  CaO,  10;  S,  20; 
Cu,  5 ;  gold  0.05  oz. ;  silver,  10.0  oz.  under  the  fol- 
lowing conditions:  (1)  furnace  capacity  175  to  200 
tons  per  24  hours;  (2)  limestone  close  at  hand;  (3) 
iron  ore  must  be  shipped  in  100  miles  by  railroad 
and  contains  no  gold  or  silver;  (4)  coke  obtainable 
at  reasonable  prices  and  of  good  quality. 

Judging  from  the  above  conditions  it  would  seem 
advisable  to  volatilize  enough  sulphur  to  make  a 
converting  grade  of  matte,  to  use  as  high  a  silica  slag 
as  is  allowable  by  the  size  of  the  furnace,  perhaps 
a  sesquisilicate,  and  substitute  lime  for  the  base  as 
much  as  possible  without  making  the  slag  too  refrac- 
tory. The  table  of  slag  formation  temperatures 
shows  that  the  lo'west  limit  for  iron  is  21.6%  without 
an  undue  rise  in  the  temperature  formation.  The 
coke  may  be  finally  reduced  on  account  of  the  sul- 
phur available  for  fuel,  but  14%  should  be  used  at 
first  and  reduced  to  the  lowest  possible  limits  with 
careful  watching. 

The  methods  of  calculation  are  given  in  the  fol- 
lowing charge  sheets  and  are  performed  in  much  the 
same  manner  as  described  in  detail  under  the  cal- 
culation of  the  lead  charge.  The  data  necessary  for 
the  computation  are  taken  from  the  table  and  matter 
previously  given. 


CHARGES  FOR  COPPER  BLAST-FURNACE        193 
CHARGE  SHEET 


Ni.-mt.  ofOt-e 


HiO  wet  o/-y    <>/, 


O*       7*e  s.nd  *\n   C*0 <ir,d  MgO       -S 
—        - "  '  -t"       %     *vt 


Of-e. 


CoA-e 


975 


*This  may  be  increased  to  1000  Ib.  by  taking  more  copper, 
probably  485  pounds. 

SLAG 

Per  cent. 

Si02 42.4 

FeO    21.6  (Fe,  16.8%) 

CaO    36.0 

S    0.8 

Cu    0.25 

COKE  (10%  ash,  1.0%  S) 
Ash.  Per  cent. 

SiO2    75.0 

FeO 10.0  . 

CaO    10.0 

MATTE 

Per  cent. 

S    25 

Cu  +  Fe  65 

S   (volatile)    80 


FACTOBS 


1*^  —  0.4  Fe  factor. 

42.4 

=  0.85  CaO  factor. 


194  TESTING  FOR  PROCESSES 

SLAG  LOSSES 

J?L  =  704  Ib.  slag,  0.25%  of  704  =  2  Ib.  Cu  slagged. 

0.424 

704  X  0.008  =   6  Ib.  S  in  slag. 
96  X  0.80   =  77    "    "  volatilized. 

83    "    "  total  loss. 

MATTE   COMPUTATIONS 

13  -4-  0.25  =  52  Ib.  of  matte. 
(22-4-52)  X  100  =  42.3%  Cu  in  matte. 
298  X  0.4  =  119  Ib.  Fe  needed. 
120    "     "     present. 

1    "    "    excess. 

AVAILABLE  IRON 

65  —  42.3%  =  22.7%  Fe  in  matte. 
52  X  0.227  =  12  Ib.  Fe  in  matte. 
298  X  0.84  =  250  Ib.  CaO  needed. 
249    "      "     present. 

1   "      "     lacking. 

These  figures  are  near  enough,  but  larger  amounts 
should  be  corrected  as  done  in  the  Lead  Charge 
Sheet. 

MATTE  FALL 

®L  X  100  =  5.3%  matte  fall. 

975 

This  is  low  and  might  well  be  increased  at  the 
expense  of  lowering  the  grade  of  matte. 

Remarks. — If  the  corrections  in  the  fluxes  amount 
to  much  in  the  first  estimate,  say  over  10  Ib.,  the 
calculations  should  be  gone  over  again.  The  charge 
may  be  increased  to  any  quantity  by  multiplying  all 
of  the  constituents  by  the  necessary  factor. 


THE  USE  OF  SLAG  DATA 


APPLICATION  OF  SLAG  DATA 


195 


Since  the  basic  problem  is  to  select  the  best  charge 
or  slag  to  smelt  a  given  ore  or  selection  of  ores, 
whether  it  is  done  for  comparison  of  two  methods  or 
to  determine  the  most  economical  slag,  all  resolve 
into  the  question,  what  mixture  put  into  the  fur- 
nace will  yield  the  greatest  net  profit. 

Example.  —  Given  a  collection  of  ores  and  fluxes 
whose  analyses  are  given  in  the  following  tables: 


CQ 
•? 

1 
o 

1 

O       jn        hj        O 

p?              £D              *C3              ^3 

w        "^        2j        $5 

P        2 

2         o         o 

"S       5-      §       § 

> 
a        ( 

o         < 

s 

p 
<-»• 

p 

1       : 

Lead  ore 

16 

10       10       60        2 

5 

Silicious  flux  or  ore 
Irony  ore    

70 
10 

4 
60 

.  .         4 

1       . 

Limestone 

1 

54 

Coke  ash,  10%  coke. 

75 

18 

9       .. 

• 

The  lead  ore  is  the  one  to  be  primarily  considered, 
it  being  the  output  of  the  mine  in  question.  The 
charges  against  this  ore  delivered  at  the  smelter  bins 
is  $2  per  ton.  The  silicious  ore  or  flux  may  be 
readily  obtained  as  a  custom  ore  and  a  treatment 
charge  of  $6  per  ton  imposed,  returning  95%  of  the 
metals  to  the  seller.  Limestone  costs  75c.  per  ton 
delivered  from  the  company's  quarry,  and  irony  ore 
costs  $4  delivered  at  the  smelter.  First  decide  if 
any  of  the  ores  should  be  roasted.  The  only  ore  with 


196  TESTING  FOR  PROCESSES 

an  appreciable  amount  of  sulphur  is  the  galena.  The 
one  very  prominent  argument  in  favor  of  not  roast- 
ing is  its  high  silver  content.  Since  one  unit  of 
sulphur  costs  25c.  to  handle,  then  in  one  ton  of 
galena  would  be  the  following'  costs  if  not  roasted : 

10  units  of  sulphur  at  25c.     $2.50  per  ton. 
Charges  if  roasted.  Per  ton. 

Roasting  H.   &  H $1.76 

Losses  Pb,  4%  of  1200  lb.,  at  Ic* 0.48 

Losses  Ag,  4%  of  50  oz.,  at  50c 1.00 

4  units  of  sulphur  left,  at  25c 1.00 

Total    $4.24 

*The  Ic.  equals  net  profit  that  would  ultimately  be  ob- 
tained if  one  pound  of  metal  were  recovered  as  market  lead. 

Difference  in  favor  of  not  roasting,  $1.74  per  ton. 
This  does  not  take  into  account  any  benefits  from 
the  use  of  a  sintered  H.  &  H.  product.  From  the 
charge-sheet  may  be  taken  the  quantities  of  ore  and 
flux  making  up  one  ton  of  blast-furnace  mixture 
and  figures  of  profit  and  loss  determined  as  follows : 

PER  TON  OF  CHARGE* 

Recovery,  Pb  93%,  Ag  95%,  Au  98%. 

CHARGES 

0.21  tons  limestone,  at  75c $0.16 

0.32  tons  irony  ore,  at  $4 1.28 

0.16  tons  lead  ore,  at  $2 0.32 

Refining  0.046  tons  base  bullion,  at  $12f 0.55 

Smelting  1  ton  charge,  at  $3 3.00 


Total  charges   $5.31 

These  figures  are  taken  from  the  lead  charge  sheet. 
tThe  cost  of  refining  is  assumed  to  cover  the  cost  of 
placing  all  the  metals  in  marketable  shape. 


DIFFERENCE  IN  SLAG  CHANGES  COST  197 

RECOVERY 

192  Ib.  Pb,  at  4c$ $7.68 

7.6  oz.  Ag,  at  50c 3.80 

Custom  ore,  0.27  ton,  at  $6 1.62 


Total  recovery    $12.10 


Profit    $6.79 

JDoes  not  include  freight  to  market. 

Next  the  charge  may  be  calculated  with  a  differ- 
ent slag  and  estimates  made  of  the  profit  with  the 
more  basic  or  acid  type.  Here  inaccuracies  creep  in 
because,  with  the  data  available,  it  is  not  possible 
to  determine  exactly  the  decreased  cost  for  smelting 
one  ton  of  charge  when  using  a  faster  running  slag 
or  to  estimate  without  actual  experiment  the  differ- 
ence in  the  recovery  of  the  metals.  The  cost  com- 
putations are  repeated  with  the  new  slag  and  the  re- 
sulting net  profit  compared  with  the  previous  re- 
sults and  deductions  drawn. 

It  will  be  noted  that  the  gold  content  of  the  sili- 
cious  ore  is  neither  charged  nor  credited  to  the  fur- 
nace charge.  This  will  introduce  a  slight  inaccuracy 
because  this  omission  is  based  on  the  assumption 
that  the  recovery  of  the  gold  will  only  be  the  95% 
which  is  returned  to  the  ore  seller.  Since  the  re- 
covery of  gold  is  given  as  98%  there  will  be  a  dif- 
ference of  3%  in  favor  of  the  recovery,  or  about  16c. 
in  this  case,  which  is  well  within  the  limits  of  errors 
for  the  example. 


COST  DATA 
GENERAL 

When  figuring  costs  and  drawing  up  designs,  it 
is  to  be  understood  the  following  figures  are  only  ap- 
proximate and  are  intended  for  use  in  making  pre- 
liminary estimates  only.  The  question  of  costs  is 
a  vital  one.  The  success  of  a  method  of  treating  an 
ore  depends  wholly  upon  dollars  and  cents  and  not 
upon  the  fact  that  the  process  in  question  is  a  scien- 
tific possibility.  In  making  final  estimates  the  entire 
scheme  must  be  carefully  analyzed  and  itemized  step 
by  step,  taking  care  to  determine  the  cost  of  each 
item  correctly.  An  underestimate  of  a  single  item 
may  cause  the  financial  failure  of  a  project  as  well 
as  ruin  the  engineer's  professional  reputation.  Cost 
data  may  be  correct  for  one  mining  camp  and  en- 
tirely wrong  for  another  a  few  miles  distant.  After 
completing  a  cost  sheet  it  is  well  to  add  from  5  to 
10%  to  the  total  as  a  factor  of  safety. 

ORE  TREATMENT 
AMALGAMATION  AND  CONCENTRATION 

STAMPS,  PLATES,  AND  VANNERS 

Per  ton.  Stamp  units. 

California    $0.40— $0.50  10—40 

Black  Hills  0.40—  0.50  200 

Alaska    0.28—  0.30  200 

Gilpin  county    0.75—  1.00  10—40 

PAN   AMALGAMATION 

Cost  per  ton  of  ore  treated  equals $2.35 


COST  OF  CONCENTRATING  MILLS 


199 


COARSE  CONCENTRATION,  JIGS,  ROLLS,  AND  TABLES 

Per  ton.  Ton  units. 

Missouri    $0.25— $0.30  100—  400 

Colorado  and   Utah 0.75—1.00  100—150 

Large  Montana  plants 0.35—  0.50  1000—2000 

Steam  stamps,  Lake  Superior.     0.25—  0.35  1000 — 2000 

COMBINATION    MILLS 

Per  ton.    Stamp  units. 
Wet  stamping,  vanners,  and  pans.  $1.50 — $5.00          60 — 100 

*COST   OF   CONCENTRATION   MILLS 

Tons 


an.  cap. 
50,000 
15,000 
75,000 


Cost  per  ton 
an.  cap. 


Coarse  concentration  of  galena, 
(Joplin)  

Mixed  sulphides  with  fine  grind- 
ing (Joplin)  

Frame  structure,  Coeur  d'Alene, 
date  of  erecting,  1900 

St.    Louis    S.    &   R.   R.   Co.,   steel 

structure,   1900    (Missouri) 300,000 

Silver  Lake  mill,  fine  grinding, 
mixed  sulphides,  hillside  site 
(San  Juan,  Col.) 75,000 

Boston  Con.,  fine  grinding,  copper 
ore,  electric  power  furnished 
(Utah)  1,000,000 

Garfield  mill,  Utah  Copper  Co.,  fine 
grinding,  erected  1907  (Utah) . .  2,200,000 

Ohio  Copper  Co.,  finished  price,  es- 
timated (Utah)  1,000,000 

*Ingalls. 

CHLORIDIZING   MILLS 

Per  ton.    Stamp  units. 
Dry  stamping,  roasting,  and  pans.  $6.00 — $7.00          60 — 150 

MAGNETIC   SEPARATION 

Per  ton   $0.25— $0.75 


$0.12— $0.16 
0.70—  0.90 
0.60 

0.80 

1.33 

1.50 
1.85 
1.50 


200  TESTING  FOR  PROCESSES 

COST  OF   MAGNETIC   SEPARATION  PLANTS 

Tons  Cost  per  ton 

ann.  cap.  ann.  cap. 

New  Jersey  Zinc  Co.   (N.  J.) 300,000  $1.75 

Fine     grinding,     Wilfley     tables, 

Wetherill  separators   15,000  3.00—  4.00 

OIL  FLOTATION 

The  cost  will  be  approximately  the  same  as  for  wet 
fine-concentration  under  the  same  conditions  and 
will  not  exceed  the  latter  by  more  than  10  to  15c. 
per  ton,  with  a  smaller  cost  of  installation  and  «will 
give  a  higher  extraction  on  certain  ores  that  are 
difficult  or  impossible  to  treat  by  wet  methods. 
CHLORINATION 

Per  ton.          Ton  units. 
Barrel  process   $3.50 — $5.00        150 — 200 

GENERAL 

BOASTING 

Per  ton. 

Heap  and  stall $0.50 — $0.75 

Hand  reverberatory    1.80 —  2.25 

Mechanical 0.35—  1.00 

Chloridizing    2.00—  2.50 

HORSE-POWER  PER  TON  OF  ORE  TREATED  PER  DAY 

Type  of  mill.                                     Mesh.  Hp. 

Stamps  and  vanners  20 — 40  0.75 — 1.0 

Coarse   concentration    10 — 20  0.5  — 0.8 

Combination  stamp   16 — 30  1.5  —1.75 

Chloridizing  stamp   (dry)    to  16  2.0  — 2.5 

Chloridizing  stamp   (wet)    to  40  4.0  — 4.5 

Magnetic  separator    0.25 — 0.5 

Cyanidation   (dry)   roll-crushing....   20 — 30  0.5  — 0.8 
Cyanidation    (wet),  stamp  crushing 

and  sliming  to  80  0.75 — 1.5 


MISCELLANEOUS  COSTS 


COST  OF  POWER  PER  HORSE-POWER  DAY 


201 


Steam-power,  non-condensing  engines,  coal  at  $4.50  =  16 
— 18  cents. 

Condensing  engines  =  14  —  16  cents. 

Compound  condensing  engines  =  11  —  14  cents. 

For  each  $1  increase  in  cost  of  coal  add  l1/^,  1%,  and  Ic. 
respectively. 

Electric-power,  $40  —  $65  per  yearly  horse-power  =  11 
— 18  cents. 

MISCELLANEOUS    ITEMS 

Stamp  batteries:  Cents. 

Wear  and  tear 7    — 12  per  ton  ore 

Labor    8    —16     "      "      " 

Power    8    — 12     "     "     " 

Coarse   crushing    (Blake   and   gyra- 
tory)         21/2—3     "      "     " 

Rolls,  dry  crushing 7  — 10  "  " 

Rolls,  wet  crushing  3  —  7  "  " 

Huntington  mills    7  — 25  "  "      " 

Tube-mills    35  —60  "  "     " 

Filter  pressing  (Moore  and  Butters)   10  — 30  "  "      " 

CYANIDATION 

Per  ton.  Ton  units. 

Dry-crushing  with  rolls  and  leach- 
ing    $1.50— $2.00  100—200 

Wet-crushing,  leaching,  and  filter- 
pressing,  or  wet-crushing, 

sliming,  and  filter-pressing...  0.85 —  2.00  100 — 200 

CYANIDE   PRACTICE,   ITEMIZED   COSTS 

Cyanide,  18  to  20c.  per  pound. 

Lime,  %c.  per  pound. 

Zinc  precipitation  =  4c.  per  ton  solution. 

Labor,  cost  for  actual  treatment. 


202  TESTING  FOR  PROCESSES 

Monthly  capacity,  Cost, 

in  tons.  cents, 

1,000 37.4 

1,500  25.8 

2,000   18.7 

2,500  15.0 

5,000  7.4 

7,500  6.0 

10,000  4.6 

15,000   3.5 

20,000  2.6 

Per  month. 

Manager $100 — $200 

Chemist 100—  125 

Foreman    80 —  100 

Maintenance  and  repairs  should  be  figured  at  from 
8  to  10%  per  annum,  calculated  on  the  original  cost 
of  the  plant.  Clean-up  and  realization  of  the  gold 
will  cost  8  to  lOc.  per  oz.  of  fine  gold.  Sundries  will 
amount  to  1  to  1%%  °^  total  cost  of  treatment.  De- 
preciation may  be  figured  at  10%  per  annum, 
of  cost  of  plant;  this  is  then  reduced  to  figures 
based  on  tonnage  by  dividing  by  total  number  of 
tons  treated  per  year.  General  charges  may  include 
a  portion  or  all  of  the  following  expenses,  which  are 
beyond  the  control  of  the  metallurgist:  (1)  general 
manager's  salary;  (2)  consulting  engineer's  salary; 
(3)  mine  office  expenses;  (4)  head  or  other  office 
expenses;  (5)  directors'  fees,  etc. 

A  modern  plant  of  steel  construction,  fully 
equipped  for  classifying  and  treating  battery  pulp, 
can  be  erected  for  $300  per  ton  treated  daily.  Fig- 
ures for  plants  of  other  design  and  construction  may 
vary  from  $180  to  $600  per  ton  treated  daily.  The 


MISCELLANEOUS  COSTS  203 

itemized  costs  include:   (1)  excavations;    (2)  foun- 
dations and  supports;   (3)   vat  precipitation  plant; 
(4)  machinery;  (5)  buildings;  (6)  tailing  disposal. 
COST  OP  HYPOSULPHITE  LIXIVIATION 

DBY  STAMPING  AND  CHLORIDIZING  BOASTING 

Crushing    $1.36 

Roasting  and  salt  (6%) 2.68 

Labor  in  leaching 0.27 

Chemicals 0.30 

Superintendence   1.02 

Heating,  lighting,  pumping,  and  repairs 0.75 

Total  cost  per  ton $5.71 

COST  OP  HUNTINGTON-HEBERLEIN  POT-ROASTING 

ASSAY    OF   ORE 

Per  cent. 

Pb 50 

Fe    15 

S    22 

SiO2    8 

A12O3,  etc 5 

Flux  used,  344  Ib.  of  limestone  and  130  Ib.  of 
quartz.  The  ore  was  80%  of  the  charge.  It  was 
planned  to  make  a  slag  consisting  of  Si02  30%, 
FeO  40,  CaO  20. 

Items.  Cost. 

Crushing  1  ton  at  lOc $0.10 

Mixing  1  ton  at  lOc 0.10 

Roasting  1  ton  at  63c 0.63 

Delivering  1.1  ton  to  converter  at  12c 0.13 

Converting  1.1  ton  at  60c 0.66 

Breaking  0.9  ton  at  60c 0.54 

Total   $2.16 

Cost  per  ton  of  ore  is  $2.16  •*•  0.80  =  $2.70. 


204  TESTING  FOR  PROCESSES 

Treatment  by  the  Salvesberg  process  will  cost  less 
by  the  amount  of  the  preliminary  roast,  which  is  63c. 
per  ton  of  charge  or  79c.  per  ton  of  ore,  making  the 
costs  respectively  $1.53  and  $1.91. 

COST  OP  LEAD  SMELTING 

AVERAGE  PER  TON   OP  CHARGE  AT  DENVER  AND  PUEBLO,  COLORADO 

Coke    $0.84 

Powder,  supplies,  labor 1.66 

General   expense    0.16 


Total    $2.66  exclusive  of  roasting 

COST  OF  SILVER-LEAD   SMELTERS 

Tons  Cost  per  ton 

ann.  cap.  ann  cap. 
Modern  blast-furnace  works;  lower 

figure  usual   330,000  $2.30—  3.00 

REFINING  LEAD  BULLION 

Parkes  Process. — Cost  =  $3  to  $5  per  ton  based  on 
actual  working  costs  to  which  must  be  added  inter- 
est, expressage,  brokerage  and  treatment  of  by-prod- 
ucts, making  the  total  approximately  $10  per  ton. 
The  following  is  an  itemized  account : 

Items.  Cost. 

Labor   $1.968 

Spelter    0.861 

Coal    0.496 

Coke   0.521 

Supplies,  repairs,  and  general  expenses...     0.289        $4.135 

Interest   1.317 

Expressage    1.085 

Parting  and  brokerage 2.121 

Re-working  by-products   1.492          6.015 

Total     $10.150 


COST  OF  COPPER  SMELTING  205 

COST  OF  PAEKES  PBOCESS  REFINERY 

Tons  base 

bullion. 
Modern  equipment  30,000        $6.66 

COST  OF  COPPER  SMELTING 

TENNESSEE  COPPER  CO.,  1903 

Per  ton. 

Coke,  0.1283  tons,  at  $4.93 $0.6082 

Quartz,  0.0958  tons,  at  90c 0.0862 

Supplies,  including  coal  for  power 0.1266 

Labor  and  superintendence   0.2111 

Total     $1.0321 

LEADVILLE,  COLORADO,  1900 

Average  cost  for  one  plant,  $3.64  per  ton  matte  shipped. 

MT.   LYELL,   TASMANIA,    1902 

Per  ton. 

Mining $0.5002 

Stripping    0.5000 

Smelting 3.3648 

Converting    0.3612 

Total    $4.7262 

This  is  the  average  cost  for  159,450  tons  ore,  average 
assay  Cu  2.36%,  Au  0.069,  Ag  2.23  oz.;  18,537  tons  fluxes, 
average  assay  Cu  1.70%,  Au  0.026,  Ag  0.24  oz.;  185,689  tons 
custom  ore;  total,  183,676  tons. 

MT.    LYELL,  1905 

Per  ton. 

Labor,  coke  fluxes,  re-treating  flue  dust $0.70 

Total  cost,  including  all  charges 1.50 

This  is  the  average  cost  for  159,450  tons  ore,  average 
ing  38  to  45%  copper,  from  ore  averaging  2.3%  copper. 


206  TESTING  FOR  PROCESSES 

COST  OF  COPPER  SMELTING  WORKS 

Cost  per 
Tons  ann.        ton  ann. 

Blast  furnace  plant,  semi-pyritic 
process,  no  roasting,  1901 330,000  $1.70 

Balaklala  works,  McDougall  roast- 
ers, blast-furnaces,  reverberatory 
furnace;  cost  includes  25c.  for 
converter  plant,  1907-08  (Cal.) .  437,500  2.25 

Washoe  plant,  cost  includes  con- 
centration and  discarded  equip- 
ment (Mont.)  3,000,000  3.58 

Highland  Boy;   total  cost  to  date 

(Utah)   300,000          3.24 

Garfleld  works;  extensive  pro- 
visions for  expansion  (Utah) . . .  800,000  7.50 

COST  OF  COPPER  LEACHING 

Henderson  Process. — Including  crushing  roasted 
pyrite  clinker,  muffle  chloridizing  roasting,  leaching 
and  precipitation,  with  common  labor  at  $1.50  per 
day,  $1.87  per  ton. 

COST  OF  ELECTROLYTIC  COPPER  REFINING 

Items.  Cost. 

Labor    $0.75 

Current    1.50 

Supplies,  acid,  coal,  etc 0.67 

Lighting 0.80 

Working  costs  $3.00 

Interest    on    permanent    capital    invested     (5%    of 

$2,000,000)    3.20 

Depreciation   (10%  of  $507,000) 1.60 

Total    $7.80 

Based  on  a  daily  treatment  of  85  tons  of  anodes. 


ELECTROLYTIC  PROCESS          207 

BETTS  ELECTROLYTIC  PROCESS 

Items.  Cost. 
Power  7.6  hp.  days  total  at  $50  per  electric  horse- 
power year   $1.06 

Tank  room,  platform,  and  repair  labor 0.86 

Melting  lead,  labor,  supplies,  and  repairs 0.38 

Coal  for  melting  lead 0.13 

Chemicals,  6  Ib.  SiF6  at  6c $0.36 

%  Ib.  glue 0.07 

0.43 

Slime  treatment,  except  power  and  assaying,  includ- 
ing parting   0.96 

$3.82 
Credit  about  20  Ib.  copper  at  3c 0.60 


Net  cost  $3.22 

Other  expenses  same  as  in  Parkes  process 6.00 

Total  cost $9.22 

COST  OF  ZINC  SMELTING 

Items.  Cost. 

Labor  (not  including  repairs  and  renewals) $6.62 

Fuel,  3  tons  at  75c.  per  ton 2.25 

Fine  coal  or  coke  for  reduction,  %  ton  per  ton  of 

ore  at  84c.  per  ton 0.47 

Clay  for  retorts,  0.1  ton  of  ore  at  $2.60  per  ton 0.26 

Repairs,  renewals,  and  sundry  supplies,  also  putting 

repaired  furnaces  in  operation  (heating  up) 0.75 

Total  cost  per  ton  blende  concentrate $10.35 

This  includes  cost  of  roasting  in  a  hand-stirred  rever- 
beratory  roaster. 


208  TESTING  FOR  PROCESSES 

COST  OF  ZINC  SMELTERS 

Cost  per 

Tons  ann.  ton  ann. 
Natural-gas     fuel     with     roasters 

(Kansas)    25,000  $8.00—10.00 

Coal  fuel,  gas-producers,  regener- 
ative furnaces  18.00 

Sulphuric   acid  works,   additional 

cost   .  5.00 —  6.00 


REFERENCES 

AMALGAMATION: 

Lodge,  'Metallurgical  Laboratory  Experiments.' 

Richards,  'Ore  Dressing,'  Vol.  I. 

Rose,    'Metallurgy    of    Gold';    Pan   Amalgamation    of 

Silver. 
Collins,  'Metallurgy  of  Silver.' 

ANALYSES: 

Lowe,  'Technical  Methods  of  Ore  Analysis.' 

Treadwell  and  Hall,  'Analytical  Chemistry,'  Vol.  II. 
Gases. 

Hempel,  'Gas  Analysis.' 

Damour,  Queneau,  'Industrial  Furnaces.' 
CHLORINATION: 

Rose,  'Metallurgy  of  Gold.' 
CYANIDATION: 

Julian   and    Smart,    'Cyaniding   of   Gold   and   Silver 
Ores.' 

Hofmann,  'Hydro-Metallurgy  of  Silver.' 

Lodge,  'Metallurgical  Laboratory  Experiments.' 

Furman,  'Manual  of  Practical  Assaying.' 

Clennell,  'Chemistry  of  Cyanide  Solutions.' 

Rose,  'Metallurgy  of  Gold.' 
CONCENTRATION  AND  ORE  DRESSING: 

Richards,  'Ore  Dressing,'  Three  Vols. 
COST  DATA: 

General  Engineering  Co.,  'Ore  Testing  Bulletin.' 

Richards,  'Ore  Dressing,'  Vol.  I  and  II. 

Austin,  'Metallurgy  of  the  Common  Metals.' 

Gillette,  'Cost  Data.' 
CHLORIDIZING  ROASTING— SILVER  LIXIVIATION: 

Hofmann,  'Hydro-Metallurgy  of  Silver.' 

Collins,  'Metallurgy  of  Silver.' 
CALORIMETRY,  PYROMETRY: 

Damour,  Queneau,  'Industrial  Furnaces.' 


210  REFERENCES 

ELECTROLYTIC  REFINING: 

Ulke,  'Modern  Electrolytic  Copper  Refining.' 

Betts,  'Lead  Refining  by  Electrolysis.' 
FURNACE  TESTING: 

Kent,  'Steam  Boiler  Economy.' 

Damour,  Queneau,  'Industrial  Furnaces.' 
GENERAL  METALLURGY: 

Austin,  'Metallurgy  of  the  Common  Metals.' 

Schnabel,  'Handbook  of  Metallurgy.'    Vol.  I,  Copper- 
Lead — Silver-Gold.     Vol.   II,   Zinc-Cadmium-Mer- 
cury -  Bismuth-Tin     Antimony-Arsenic-Nickel-Co- 
balt-Platinum-Aluminium. 
OXIDIZING  ROASTING: 

Peters,  'Principles  of  Copper  Smelting.' 
PARKES  PROCESS: 

Hofman,   'The  Metallurgy   of  Lead  and   Desilveriza- 
tion  of  Base  Bullion.' 

Collins,  'Metallurgy  of  Lead.' 
POT  ROASTING: 

Ingalls,  'Lead  Smelting  and  Refining.' 

SULPHATING  ROAST— ZIERVOGEL  PROCESS: 

Collins,  'Metallurgy  of  Silver.' 

Hofmann,  'Hydro-Metallurgy  of  Silver.' 
RETORTING  AND  REFINING: 

Richards,  'Ore  Dressing,'  Vol.  I. 

Julian  and  Smart,  'Cyaniding  of  Gold  and  Silver  Ores.' 

Rose,  'Metallurgy  of  Gold.' 
SMELTING: 

Peters,  'Principles  of  Copper  Smelting.' 

Rickard,  'Pyrite  Smelting.' 

Hixon,  'Lead  and  Copper  Smelting  and  Copper  Con- 
verting.' 

Ingalls,  'Lead  Smelting  and  Refining.' 
ZINC  ROASTING  AND  REFINING: 

Ingalls,  'Metallurgy  of  Zinc  and  Cadmium.' 


INDEX 


Page. 
Acid,    Consumption    of....      95 

Excess    of    22 

Acidity,   Total    27 

Addition  of  Chemicals 73 

Agitation 27 

Agitator    and   Leaf    Filter, 

Slime     63 

For  Sliming   43 

Alkali   Solutions    32 

Alumina   178 

In    Slags    168 

Aluminum,    Salts   of    45 

Amalgam,  Preparation  for 

Retorting     137 

Amalgamation   9 

And  Concentration 198 

For  Silver  Ores 71 

Test    for    9 

Test   for  Pan    74 

Amalgamator  and  Grinder     72 
Amount   of   Zinc   Used....      58 
Analysis  of  Cyanide  Solu- 
tion          49 

Antimonial    Ores     73 

Antimony  and  Arsenic   .  .  .    178 
In   Electrolytic   Refining 

100,  106 

Sulphide     47 

Application  of  Slag  Data.    195 
Arsenic   and   Antimony    .  .    178 

In  Cyanide  Process    47 

In   Electrolytic   Refining 

100,  106 
Arsenical   Ores    73 

Barrell  Process  18 

Barrels,  Charging  of 22 

Baryta  «.  .  .  179 

In  Slags 168 

Base  Bullion,  Amount  of.  170 

Betts  Electrolytic  Process  207 

Process  104 

Bismuth  in  Electrolytic 


Page. 

Refining    100,  106 

Blast-Furnace  Charge,  Cal- 
culations  of    192 

Charge,    Calculations    of 

a    Silver-Lead    ... 186 

Boiling   Points,   Determin- 
ation of   133 

Box  Operations,  Zinc 58 

Bullion,    Calculation    of...  189 

Refining 140 

Refining  Lead    204 

Bunsen  Burner    75 

Butters   Process    60 

Calcium   Carbonate    45 

Sulphate 76 

Carbonates,      Soluble     and 

Insoluble 46 

Carmichael-Bradford   Pro- 
cess      76 

Causes  for  Non-Extraction  50 

Caustic  Lime    45 

Charcoal,   Precipitation   of 

Gold   by    20 

Charge,      Calculations      of 

Copper    Blast-Furnace  192 
Calculations     of     Silver- 
Lead    Blast-Furnace..  186 

Leaching  of    22 

Charges,  Cost  of 196 

Estimation  of 190 

Chemical  Reactions   74 

Tests   26 

Chemicals,   Addition   of    . .  73 
Chloride    Ores,    Free-Mill- 
ing    73 

Chloridizing,         Principles 

and  Reactions    66 

Roast  of  Silver  Ores   ...  66 

Chlorination    15 

Cost  of   200 

Test 21 

Classifier,    Hydraulic    11 


212 


INDEX 


Page. 

Coarse  Gold    9 

Cobalt  in  Electrolytic  Re- 
fining      100 

Color  of  Precipitate 58 

Concentrate,  Percentage  of  9 

Treatment  of   61 

Concentrating    Mills,    Cost 

of    199 

Concentration    and    Amal- 
gamation      198 

Crushing   in    61 

Maximum  Size  of  Parti- 
cles      145 

Ratios,  Mattes   and 181 

Tests   143 

Wet    146 

Consumption,      Causes      of 

Cyanide    -.  45 

Of   Acid    95 

Of  Cyanide   26 

Of   Zinc    57 

Copper        Blast   -   Furnace 
Charge,       Calculations 

of 192 

Current    Used    in    Refin- 
ing    103 

Electrolytic  Refining  of.  99 
In   Electrolytic   Refining 

100,  106 

In  Solution    59 

Leaching,   Cost   of    206 

Matte,         Sulphatizing 

Roasting    86 

Ores,  Lixiviation  of   ....  94 
Refining,    Cost    of    Elec- 
trolytic    206 

Refining,   Efficiency  in..  104 

Slags 173 

Smelting,  Cost  of   205 

Smelting  Properties  of.  .  164 

.     Smelting  Works,  Cost  of  206 
Sulphate,  Decomposition 

of    20 

Treated  by  Cyanide   ....  46 

Cost  Data    198 

Of  B  e  1 1  s      Electrolytic 

Process 207 

Of  Charges   196 


Page. 

Of  Chlorination 200 

Of  Concentrating    Mills.  199 

Of  Copper  Leaching 206 

Of  Copper    Smelting    ...  205 
Of  Copper     Smelting 

Works 206 

Of   Cyanidation    201 

Of  Electrolytic      Copper 

Refining 206 

Of  Henderson   Process.  .  206 
Of  Huntington  -  Heber- 

lein    Pot-Roasting    .  .  .  203 
Of  Hyposulphite    Lixivi- 

tion    203 

Of  Lead    Smelting    204 

Of  Oil  Flotation    200 

Of  Parkes    Process    ....  204 

Of  Zinc   Smelters    208 

Of  Zinc    Smelting     207 

Statement     of     23 

Costs,    Miscellaneous     ....  201 

Crucible  for  Slag  Test 161 

Crushing  in  Concentration  61 
Current   Density   in   Refin- 
ing     100,  107 

Used  in  Copper  Refining  103 

Cyanidation     24 

Cost  of   201 

Double    Treatment    of.  .  55 

Of    Silver    Ores    64 

Cyanide      Consumption, 

Causes    of .  45 

Consumption    of    26 

Plant,     Laboratory     ....  54 

Poisoning     51 

Process,    Reactions    in..  51 

Solution,     Analysis     of..  49 
Solutions,     Solubility    of 

Metals  and  Minerals..  29 

in    29 

Test    on    Large    Lots    of 

Ore     53 

Cyaniding    Slime     61 

Dead    Roast    20 

Decantation   Test    44 

Desilverization   Process    .  .  80 

Dissolution,   Rate  of    33 


INDEX 


213 


Page. 
Double  Treatment  of  Cya- 

nidation    55 

Effect  of  Strong  and  Weak 

Solutions    28 

Electrical  Precipitation...  59 
Electrolysis,  Solution  for.  101 
Electrolyte,  Acidity  of.  ..  108 

Making   up    101 

Preparation  of 106 

Purification    of    101 

Electrolytic  Copper  Refin- 
ing,  Cost   of    206 

Process,  Cost  of  Betts..    207 

Processes   97 

Electromotive  Force    99 

Electroplating    60 

Electrostatic    Separation..    151 

Ellsner's   Equation    51,     52 

Elmore  Process   153 

Elutriation  Apparatus.  40,  41 
Estimation  of  Charges...  190 

Experimental   Work    77 

Extractions,  Diagram  of.  .      34 

Faraday's    Law     97 

Ferrous   Hydrate    46 

Sulphate    19,     46 

Filter  and  Slime  Agitator, 

Leaf   63 

Vacuum  Leaf  Slime   ....      62 

Fine   Concentration    148 

Fineness  of  Grinding  ....  32 
Flotation,  Cost  of  Oil 200 

Process     153 

Flow  of  Solution,  Rate   of     58 

Flue  Dust    173 

Free-Milling     9 

Chloride  Ores 73 

Test    10 

Freezing  Point   of  Metals, 

Determination   of    ....    134 
Fuel,  Calorific  Power  of..    125 

Quantity   of    170 

Furnace   Data    169 

Laboratory   Roasting    .  .      17 

Temperature    of    20 

Galena    .    179 


Page. 

Gold,  Appearance  of    25 

In   Electrolytic   Refining  100 

Pan,  Miner's   11 

Precipitation   of    19 

Smelting  Properties  of.  .  164 

Solubility  of   15 

Solution   of    18 

Spongy     25 

Goutal   Formula    126 

Grinder  and  Pan  Amalga- 
mator      72 

Grinding,  Fineness  of   ....  32 

Heat    Balance    Sheet 118 

Henderson  Process 95 

Process,   Cost  of    206 

Hoffman     178 

Hoffman's  Slag  Table 184 

Horse-Power    per    Ton    of 

Ore  Treated    200 

Huntington-Heberlein  Pot- 
Roasting,  Cost  of 203 

Hydraulic    Classifier    11 

Hydrogen   Sulphide    19 

Sulphide,       Precipitation 

by     22 

Hyposulphite     Lixiviation, 

Cost  of   203 

Silver  Salts  Soluble  in..      69 

Impurities,     Effect     of     in 

Electrolytic    Refining.    100 

Ingalls,  W.  R 76 

Insoluble  Carbonates 46 

Interference      of      Various 

Substances 71 

Investigations,      Prelimin- 
ary         24 

Iron  and  Manganese   179 

In   Electrolytic    Refining 

100,  106 

In  Slags   170 

Sulphate,    Decomposition     20 
To  Silica,  Ratio  of 187 

Jigging  Tests 147 

Laboratory   Cyanide   Plant     54 

Roasting  Furnace 17 

Leaching   and    Percolation     35 


214 


INDEX 


Page. 
Cost  of  Copper 206 

Lead,  Amount  in  Charge..    170 

Bullion,  Refining 204 

In   Electrolytic   Refining 

100,   104 

Roasting    173 

Slag,  Types  of 165 

Smelting,  Cost  of 204 

Smelting  Properties  of.  .    164 

Leaf  Filter  and  Slime  Agi- 
tator        63 

Slime  Filter,  Vacuum...      62 

LeChatelier  Pyrometer   ...    128 

Lime,  Amount   of 44 

And  Magnesia   180 

In  Cyanidation 27 

In  Slags   168 

Roasting    76 

Limestone  to  Silica,  Ratio 

of    187 

Liquator  for  Zinc  Crusts.  .      84 

Lixiviation,  Cost  of  Hypo- 
sulphite        203 

Of  Copper  Ores   94 

Tests  of   94 

Loss  in  Roasting 68 

Of  Heat    118 

Of  Weight  in  Roasting.    183 

Losses  in  Lead  Smelting..    172 
Mechanical    21 

Maclaurin    29 

Magnesia   and   Lime    180 

In  Slags 168 

Magnesium,    Salts    of    45 

Magnetic  Separation    150 

Malher  Formula    127 

Manganese  and  Iron 179 

Matte  Content   171 

Fall    170 

Mattes   and    Concentration 

Ratios 181 

Melting    Point    of    Metals, 

Determination  of    134 

Mercury    49 

Microscopic    Examination.  25 

Miscellaneous  Costs 201 

Mispickel   47 


Page. 

Moisture    in    Leaching    ...  36 
Multiple  System   of  Refin- 
ing     101 

Nickel  in  Electrolytic  Re- 
fining    100,  106 

Nitro-Prussides   49 

Non-Extraction,  Causes  for  50 

Oil  Flotation  153 

Flotation,  Cost  of  200 

Ore,  Free-Milling  9 

Testing 67 

Treatment  . 198 

Value  of  9 

Ores  Non-Amalgamating.  .  15 

Pan      Amalgamation       for 

Silver  Ores    71 

Amalgamation,   Test  for  74 
Amalgamator  and  Grind- 
er      72 

Parkes  Process    80 

Process,   Cost  of    204 

Percentage  of  Salt 66 

Percolation    and   Leaching  35 

Rate  of    35,  37 

Test  of   38 

Peters,   E.   D 182 

Platinum     in     Electrolytic 

Refining    100 

Plattner  Process   18 

Poisoning,   Cyanide    51 

Porosity  and  Rate  of  Per- 
colation    38,  39 

Pot    Roaster     78 

Roasting    76 

Roasting,   Cost   of   Hun- 

tington-Heberlein     ...  203 

Potash   in   Slags    168 

Precipitate,  Color  of 58 

Precipitation     19 

Electrical     59 

Of  Zinc 56 

Preparation      for     Sliming 

Tests    62 

Principles    of    Chloridizing  66 

Process,  Butters    60 

Siemens  &  Halske   60 


INDEX 


215 


Page. 

Processes,  Electrolytic   ...  97 

Pulverizing   32 

Purification      of      Electro- 
lyte      101 

Pyrite    46 

Pyrometer,        Standardiza- 
tion  of    131 

Pyrometry     128 

Rate  of  Dissolution   33 

Of  Flow  of  Solution 58 

Ratio,  Mattes  and  Concen- 
tration    181 

Of    Iron    and    Limestone 

to  Silica 187 

Reactions,    Chemical    74 

In  Cyanide  Process    ....  51 

Of  Chloridizing 66 

Reducing  Factors 88 

Refining  Bullion    140 

Lead  Bullion 204 

Multiple  System  of    101 

Series,    System    of 101 

Retort     -. 138 

Preparation    of    137 

Retorting     137 

Rio  Tinto  Method 94 

Roast,  Efficiency  of Ill 

Of    Silver    Ores,    Chlori- 
dizing      66 

Oxidizing    16 

Roaster,  Pot 78 

Roasting,     Cost     of     Hun- 

tington-Heberlein   ....  203 

Furnace,    Laboratory    .  .  17 

Lime    76 

Loss  in   68 

Loss  of  Weight  in   183 

Of    Copper    Matte,     Sul- 

phatizing   86 

Pot   76 

Requirements  for 16 

Rusty  Gold   16 

Salt,  Percentage  of 66 

Salts     of     Aluminum     and 

Magnesium    45 

Soluble    69 

Salvesberg   76 


Page. 

Process 204 

Sampling  Siphon 44 

Screen    Sizing   Tests    145 

Selenium     in     Electrolytic 

Refining   106 

Series  Systems  of  Refining  101 

Siemens  &  Halske  Process  60 
Silica,    Ratio    of    Iron    and 

Limestone  to    187 

Silver  as  Sulphate 69 

Chemically    Com  b  i  n  e  d 

with   Gold    25 

In    Electrolytic   Refining 

100,  106 
Lead    Blast  -Furnace 

Charge,  Calculation  of  186 
Ores,  Chloridizing  Roast 

of    66 

Ores,  Cyanidation    of    .  .  64 
Ores,  P  a,  n      Amalgama- 
tion   for    71 

Salts    Soluble    in    Hypo- 
sulphite      69 

Salts   Soluble  in  Water.  69 

Smelting  Properties  of..  164 

Siphon,   Sampling    44 

Sizing  Test   40 

Slag,   Molten    162 

Constituents,  Effect  of..  168 

Constituents,  Percentage  167 

Crucible  for  Test 161 

Data,  Application  of   ...  195 

Necessary  Superheat   .  . .  162 

Physical  Properties  of..  160 

Report   on    162 

Slags,  Copper   173 

Selection  of 175 

Types  of  Lead   165 

Slime   Agitator    43 

'Agitator  and  Leaf  Filter  63 

Cyaniding    61 

Filter,   Vacuum   Leaf    . .  62 

Recovery  of   104 

Sliming  Tests    42,  62 

Smelter  Rates    156 

Smelting,  Cost  of  Copper.  205 

Cost  of  Lead 204 

Cost  of  Zinc   .                     .  207 


216 


INDEX 


Page. 
Works,  Cost  of  Copper.  .    206 

Soda   in   Slags    168 

Solubility  of  Gold  and  Sil- 
ver         31 

Of   Metals   and   Minerals 
in    Cyanide    Solutions.      29 

Soluble   Carbonates    46 

Salts    69 

Silver  Salts  in  Hyposul- 
phite          69 

Silver  Salts  in  Water.  .  .      69 

Solution,  Copper  in   59 

For  Electrolysis    101 

Strength  of    28,     59 

Spar,  Heavy   179 

Speiss 172 

Stibnite     47 

Strength   of  Solution    ..28,     59 
Stripping  Plates,  Prepara- 
tion  of    103 

Sulphate,   Calcium    76 

Silver   as    69 

Sulphatizing    Roasting    of 

Copper    Matte    86 

Sulphide,  Antimony   47 

Sulphides,  Precipitation  of 

Gold  by 20 

Sulphur 180 

In  Ores    16,     66 

Tailing,  Value  after  Amal- 
gamation            9 

Tellurium 48 

In   Electrolytic   Refining  106 

Temperature    20,   136 

Test    for    Pan    Amalgama- 
tion          74 

For  Sliming    42 

Free  Milling 10 

Of  Percolation   38 

On  Large  Lot  of  Ore,  Cy- 
anide     ~  53 

Testing  an  Ore    67 

Porosity      of      Leaching 

Material    39 

Tests,  Chemical 26 


Page. 

For  Decantation    44 

Of    Lixiviation     94 

Of  Suitability  of  Ore  for 

Smelting     155 

Sliming   62 

Thermal  Capacity  of  Gases  119 
Thermo-Electric     Pyrome- 
ter     129,  130 

Time  Factor  in  Solution..  31 

Of  Cyanidation   53 

Of  Treatment    33 

Tin    in    Electrolytic   Refin- 
ing      106 

Titration    27 

Treatment    of   Concentrate  61 

Of   Ore    198 

Time   of   33 

Vacuum  Leaf  Slime  Filter  62 

Velocity  in   Percolation...  36 

Weight   in   Roasting,   Loss 

of    183 

Wet-Sizing    40 

Ziervogel  Process 86 

Zinc   181 

Amount  Used   58 

Box  Operations   58 

Consumption  of 57 

Crusts,  Liquator  for    ...  84 

Distillation  of    110 

In    Electrolytic   Refining 

100,  106 

In  Slags   168 

Precipitation    by    Means 

of    56 

Roasting,  Impurities  in  .  109 

Roasting,  Reactions  in    .  110 

Smelters,  Cost  of 208 

Smelting 109 

Smelting,  Cost   of    207 

Smelting,  Diagram    for 

Furnace    Test    116 

Smelting,  Furnace    Test.  112 

Smelting,  Report    on    ...  117 

Smelting,  Test   for    Ill 


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